THE 


Hydro-electrolytic 

Treatment 

of 

Copper Ores 


By 

R. R. GOODRICH 




















Hydro-electrolytic 
T reatment 

of 

Copper Ores 


By 

ROBERT RHEA GOODRICH 


Submitted in Partial Fulfilment of the Requirements for the Degree of 
Doctor of Philosophy, in the Faculty of Pure Science, 

Columbia University. 




1915 






This research was partly done in the non-ferrous laboratory 
of the Department of Metallurgy of Columbia University, under 
the direction of Dr. Edward F. Kern, and completed elsewhere. 
Acknowledgement is due to Professor Arthur L. Walker, Dr. 
Edward F. Kern and Dr. William Campbell of the Department 
of Metallurgy, for their kind advice and for the inspiration 
derived from their instruction. 


Gift 

The. University 
OCT 27 I9JS 





v3 


*i) 


1 


>o 


A paper presented at the Twenty-fifth Gen¬ 
eral Meeting of the American Electro¬ 
chemical Society, in New York City, 
April 18, 1914. 


CTO 


THE HYDRO-ELECTROLYTIC TREATMENT OF COPPER ORES 

By Robert Rhea Goodrich. 


It is the intention of this paper to give a brief summary of 
the practised and suggested hydrometallurgical processes for the 
extraction of copper from its ores. The review of the literature 
as given contains the salient points of what has been done in the 
history of the subject. 

The problems of economically treating low-grade copper ores 
have turned the attention of metallurgists toward such hydro- 
metallurgical problems as are encountered with low-grade sili¬ 
ceous, oxidized and sulphide ores, and concentrates, especially 
where water-power is cheap and fuel is expensive. The treat¬ 
ment of tailings from concentrates is another promising field 
for leaching methods, there being no other known possible method 
for economically extracting their copper contents. The same is 
also true of complex refractory ores which are not amenable to 
smelting for the recovery of their several metals. And, lastly, 
there are metallurgists who entertain hopes of discovering a 
leaching method which will radically change all methods of copper 
extraction to something quicker and cheaper than the present 
smelting methods. 

Classification of Hydrometallurgical Processes. 

I. Purely Chemical Methods: 

1. Alkali processes. 

2. Sulphite “ 

3. Sulphate “ 

4. Chloride ** 

Copper is dissolved and precipitated by chemical reagents- 

II. Electrolytic Methods: 

/. Sulphate processes. 

2. Chloride “ 


207 


208 


ROBERT RHEA GOODRICH. 


Copper is dissolved chemically and is precipitated electro- 
lytically. The deposition is usually accompanied by regeneration 
of the solvent. 

All acids react more or less with the constituents of the ore, 
causing: 

a. Consumption of acid. 

b. The bringing in of elements detrimental to the process. 

Iron, arsenic, antimony and bismuth, while detrimental, are not 
necessarily fatal to an acid process. If lime, magnesia, zinc or 
manganese occur in large quantities in the ore, acid processes 
are not applicable—the limit can only be determined by experi¬ 
ment. While calcium carbonate is detrimental, calcium sulphate 
is not. Alumina is undesirable, but not necessarily very injurious. 

Many oxidized ores are improved by roasting. 

All sulphide ores require roasting for most of the leaching 
processes, the exception being chalcocite ores, which may be 
leached direct. (Greenawalt, ‘‘Hydrometallurgy of Copper,” 1912 
Ed., Chap. IX.) 

I. Purely Chemical Methods: 

1. Alkali Processes. 

The alkali processes have not met with much encouragement 
in the hydrometallurgical extraction of copper from its ores, due 
largely to the low and slow solubility of copper minerals in 
solutions of the alkalies. Ammonia and ammonium compounds 
are the only alkaline solvents tried for the leaching of copper 
oxide ores on a commercial scale. (Greenawalt, p. 172.) 

The Mosher-Ludlow Ammonia-cyanide Process. 

This process is applicable to ores containing oxide and car¬ 
bonate of copper. At the ordinary temperature, ammonia (NH 3 ) 
forms a stable compound [Cu(NH 3 ) 2 ] which readily dissolves 
in water containing slight excess of ammonia. At the boiling 
point of the solution the compound is broken up, hydrated copper 
oxide being precipitated. The distilled ammonia may be con¬ 
densed and used for further treatment of the ore. When the 
ore contains gold and silver as well as copper, a weak solution 
of potassium cyanide is used to extract the precious metals 


TREATMENT OF COPPER ORE. 


209 


subsequently to the copper. The gold and silver are precipitated 
by zinc. (Electrochemical and Metallurgical Industry, Mar., 
1908, p. 128; Greenawalt, p. 172.) 

2. Sulphite Processes. 

The Neill Process is applicable to ores containing oxide and 
carbonate of copper. When cupric oxide is treated with water 
through which sulphur dioxide is passed, the following reaction 
takes place: 3C11O + 3S0 2 = 3CUSO3. The cupric sulphite 
formed is insoluble in water, but is readily soluble in water con¬ 
taining an excess of sulphur dioxide. This constitutes the 
leaching step. Then the clear liquid is drawn from the ore and 
heated to drive off the excess sulphur dioxide, when a bright 
red, heavy precipitate [Cu 2 S 0 3 .CuS 0 3 ] is thrown down. Next, 
the liquid is run over iron to precipitate any copper that may 
remain in solution as sulphate. The liquid which contains no 
copper finally goes to waste. The experimental plant of the 
Montana Ore Purchasing Co., at Butte, was put up by Neill. 
Washing was difficult on account of ferric oxide which formed 
in the leaching charge. The results were: Ore, 3.15 percent Cu ; 
tailings, 0.31 percent Cu; extraction 90 percent. (Greenawalt, 

p- 178-) 

The Van Arsdale Process consists in precipitating the copper 
from cupric sulphate solution by adding sulphur dioxide (3CuS0 4 
-j- 3S0 2 -f- 4H 2 0 = [Cu 2 S 0 3 .CuS 0 3 ] -f- 4H 2 S0 4 ), and heating 
with or without pressure (Cu 2 S 0 3 .CuS 0 3 -j- 4H 2 S0 4 = Cu + 
2 CuS 0 4 + 2H 2 S0 4 + 2S0 2 + 2 H 2 0 ), thereby simultaneously 
regenerating acid solution for leaching the ore. The regenerated 
sulphuric acid is double the amount required to dissolve the 
amount of copper precipitated. Only a part of the copper is 
precipitated. The process being cyclic, there is no particular 
harm in returning to the ore a solution containing a considerable 
amount of unprecipitated copper sulphate. (E. and M. J., June, 
1908; U. S. Patent, Mar. 31, 1903, No. 723,949; Greenawalt, 

P- 1 79 -) 

3. Sulphate Processes. 

Oxidized ores containing copper as oxide or carbonate may be 
treated directly with dilute sulphuric acid solutions. Sulphide 
ores, with the possible exception of certain chalcocite ores, should 


210 


ROBERT RHEA GOODRICH. 


be roasted prior to leaching. The copper may be dissolved as 
sulphate either by' (i) sulphuric acid or (2) metal sulphates. 
Reactions during leaching with sulphuric acid are: CuO + * 
H 2 S 0 4 = CuS 0 4 + I~f- 2 0 , and during precipitation: CuS 0 4 
Fe = Cu + FeS 0 4 . Theoretically, 1.68 lb. 66° B. sulphuric 
acid and 0.88 lb. iron are required to produce 1 lb. of copper. 

Copper has been dissolved from its ores by ferric sulphate 
solutions containing free sulphuric acid: xH 2 S 0 4 + Cu 2 S + 
2Fe 2 (S0 4 ) 3 = 2 CuS0 4 + 4FeS0 4 + S + xH 2 S 0 4 . Cuprous 
sulphide is slowly acted upon by solution of ferric sulphate. 
Most of the improvements of this simple process are based on 
the regeneration of the solvent. (Greenawalt, p. 180.) 

Sulphuric Acid Leaching of Oxidized Copper Ores 

at Clifton, Arizona. 

The Arizona Copper Company has been leaching oxidized 
surface ores at Clifton on a large scale since 1893. 

The sulphuric acid is manufactured from roaster gases. The 
Joy Mine, Morenci, furnishes the pyrite ore, containing 1.5 per¬ 
cent copper, which is crushed to 2-inch. The fines are roasted 
in a Herreshoff five-deck furnace. The coarse material goes to 
lump burners. The cinder from the roasting furnaces is smelted 
in blast furnaces. The resulting acid (chamber acid) is 52 0 B. 

The ore which contains 2.5 percent of copper is conveyed from 
the mines at Metcalf to the “Oxide Mill” at Clifton. The copper 
exists chiefly as malachite, together with some sulphides. By 
wet concentration, 25 percent of the copper is extracted as con¬ 
centrates, carrying 7 to 10 percent of copper, which are smelted. 
Slimes carrying 2.4 percent of copper go to' the slime pond for 
settling; they contain too much soluble alumina to leach. The 
tailings, size 1 inch to sand—75 percent of them larger than l /% 
inch—go to the leaching vats. Circular wooden vats, with per¬ 
forated false bottom for filter, are used for leaching. The solu¬ 
tion is circulated by centrifugal pumps, one being supplied to 
each tank. Successive solutions are applied. The first solution 
to fresh ore (concentrator tailings) is one high in copper and 
low in acid. This solution, leaving the vat with practically no 
free acid, passes on to the precipitation tanks. There is then 
applied to the ore a solution containing more free acid and less 


TREATMENT OF COPPER ORE. 


2 I I 


copper, followed by a fresh acid solution and lastly a water wash. 
1 he copper contents of the liquors is precipitated on scrap iron. 
There being no free sulphuric acid, the consumption of iron is a 
minimum. Liquors free from copper are run to waste in earthen 
reservoirs and sink in the ground, thereby preventing contamina¬ 
tion of the streams. There are consumed 2.6 lb. of 52 0 B. acid 
{1.82 lb. 66° B. Com.) per pound of copper extracted. No 
general rule can be given for strength of acid. The more soluble 
the gangue, the weaker should be the acid. The ore carries a 
trace of gold, which apparently is lost. (Greenawalt, p. 183.) 

Leaching Plant at the Snowstorm Mine. 

The vein at Larson, Idaho, is a replacement in quartzite. Only 
the oxidized ore of the upper workings (consisting of cuprite, 
malachite and chrysocolla) is treated. The copper content is 3 
percent, with small gold and silver values. The process is : Crush 
the ore and treat in agitators with bleaching powder and sul¬ 
phuric acid, thus converting the copper, gold and silver into 
chlorides. Separate the liquor, and precipitate the copper and 
gold on scrap iron. Treat the residue with hyposulphite of 
soda to dissolve the silver chloride, which subsequently precipi¬ 
tate by sodium sulphide. The saving is 90 percent of assay value 
of the ore. (Greenawalt, p. 187.) 

Copper Leaching Plant at the Gumcshevsky Mine, Russia. 

The mine was shut down in 1871. The mine owners con¬ 
tracted with an acid manufacturer to manufacture acid on the 
property from pyrite carrying 3.5 to 8 percent of copper; to sell 
53 0 B. acid to them at $4.32 per ton; to extract the copper from 
the burned pyrite, leaving less than 0.3 percent of copper in the 
tailings; to pay the owners $145 to $175 per ton of copper pro¬ 
duced. The method used to extract the copper from the burned 
pyrite was: Roast burned pyrite with addition of sulphuric acid 
in a muffle furnace at 450° C. to 550° C., bringing copper into 
soluble condition; leach with water; leach with barren solution 
after precipitating on cast iron plates; leach with dilute sul¬ 
phuric acid; precipitate copper at boiling temperature on cast 
iron plates. Consumption of acid, 2 lb. (66° B. Com.), and of 
cast iron, 1 to 2 lb. per pound of copper produced. 


212 


ROBERT RHEA GOODRICH. 


The mine owners worked the extensive dump left from the 
earlier working of the mine, when it was worked for oxidized 
ore only. The process was: Crush dump material by rock 
breaker and Chilean mills; leach with dilute sulphuric acid; 
precipitate copper by cast iron, granulated or in plates. The 
oxidized dump material treated contained 0.75 percent copper; 
of this 0.43 percent was recovered, the balance going into the 
tailings. Extraction, 57 percent. Insoluble copper existed as 
silicate and native copper. Consumption of acid, 7.4 lb., and of 
iron, 1.9 lb. for 1 pound of copper produced. Thus only 23 per¬ 
cent of acid consumed was usefully employed in dissolving cop¬ 
per ; the balance acted on worthless gangue. All tanks, both for 
leaching and for precipitating, were rectangular and made of 
concrete. The agitator moved longitudinally in the leaching 
tank. (Greenawalt, p. 187; Inst, of Min. and Met. Bull., No. 65; 
Trans. I. M. M., XIX, p. 212; Min. Ind., 1910, p. 210.) 

The Laist Process. 

The process is based on the use of sulphuric acid as the solvent 
and hydrogen sulphide as the precipitant. The steps in the 
process are as follows: 

1. Dissolving copper with dilute sulphuric acid: CuO -f- 
H 2 S 0 4 = CuS 0 4 + h 2 o. 

2. Precipitating copper by hydrogen sulphide and regenerat¬ 
ing sulphuric acid: CuS 0 4 + H 2 S = H 2 S 0 4 -j- CuS. 

3. (a) Manufacturing hydrogen sulphide by reducing calcium 
sulphate (gypsum) with coal, which takes place at 1800° F. 
(982° C.) : CaS 0 4 + 4 C = CaS + 4CO. 

(b) Decomposing calcium sulphide by carbon dioxide and 
water to hydrogen sulphide and calcium carbonate: CaS + H 2 0 
+ C 0 2 = H 2 S + CaC 0 3 . 

4. Converting copper sulphide precipitate to metallic copper 
by roasting and reducing in furnaces. In practice there was 
used for reaction 3 (a) about 2^ lb. gypsum and ij4 lb- of 
coal, which is but slightly above the theoretical amount. (Greena¬ 
walt, p. 204.) 


TREATMENT OF COPPER ORE. 


213 


♦ 


Roasting and Leaching Tailings at Anaconda, Montana. 

During the summer of 1912 experiments were made on regular 
concentrator tailings by roasting and leaching. The results were 
so satisfactory that it was decided to continue the work on a 
larger scale. In February, 1913, construction was begun on an 
8o-ton roasting and leaching plant. It was decided to precipitate 
copper by scrap iron and to experiment with precipitation by elec¬ 
trolysis and by hydrogen sulphide. 

The roasting experiment was conducted in a regular No. 64 
improved McDougal furnace. Two fire boxes were placed so that 
the flame could enter the third or the fourth floor (from the 
top). The third floor was selected as the better. The best 
results were obtained by oxy-chloride roasting. The firing was 
regulated by a pyrometer inserted on the fourth floor. The 
temperature was held at iooo° F. (538° C.). On the upper three 
hearths the ore was partly roasted. One percent by weight of 
sodium chloride was added on the fourth floor. 

The leaching was by percolation: No. 1 solution contained 
3.5 percent of H 2 S 0 4 and 10 percent of NaCl; No. 2 solution 
contained 6 percent of H 2 S 0 4 and 10 percent of NaCl. Then 
followed water washes. The precipitation was done by hydrogen 
sulphide. No. 1 solution was the only one precipitated. No. 2 
solution became No. 1 solution on the next vat. The acid 
strength was brought up in the precipitating vat. It was not 
necessary to add salt (NaCl), as the solution picked it up from 
the roasted ore. 

Summary of two months operation: 

In calcines to leaching plant: 10.4 lb. copper per ton, 0.46 oz. 
silver per ton. 

Percentage of recoverable copper, 85.4 percent; percentage of 
recoverable silver, 91.1 percent. 

Conclusions: 

It is not advisable to leach high-grade 3 percent copper material, 
which may be concentrated. 

Material should be crushed to 15-mesh preparatory to roasting 
and leaching. 


214 


ROBERT RHEA GOODRICH. 


Copper set free by crushing should be taken out as high-grade 
concentrate and smelted. 

Sulphuric acid can be cheaply made at a smelting plant. 

It is estimated that copper can be produced for 6.5 cents per 
pound. This is possible because tailings are already mined, 
crushed and sized; cheap acid may be made from roaster gases; 
operations are on a large scale. (M. and E. W., Yol. 39, No. 13, 
p. 545, Sept. 27, 1913; No. 14, p. 599, Oct. 4, 1913; Frederick 
Eaist.) 


Experiments with Ferric Sulphate at Cananea. 

The dissolving of cuprous sulphide from the ores, by means 
of ferric sulphate solution, is in accordance with the reaction: 
Cu 2 S -j- 2Fe 2 ,(S0 4 ) 3 = 2 CuS0 4 + 4FeS0 4 + S'. Thus, for 
1 lb. of copper going into solution, 6.3 lb. of anhydrous ferric 
sulphate is reduced. Copper oxide is dissolved by ferric sul¬ 
phate solutions, as expressed by: 3C11O + Fe 2 (S 0 4 ) 3 = 3CuS0 4 
+ Fe 2 0 3 . Thus in dissolving copper oxide, 2.1 lb. anhydrous 
ferric sulphate is consumed for 1 lb. copper going into solution. 

At Cananea the consumption was found to be 4.37 lb. ferric 
sulphate per pound of copper extracted. The content of solution 
in ferric sulphate was of small import, provided basic iron salts 
were absent from it. The solution worked with as small an 
amount as 1 percent. The extraction with a 2 percent solution 
was nearly as complete as with a 7 percent solution. 4 From 
leaching, the liquor went to the precipitation tank, in which 
iron was used for precipitating the copper according to : CuS 0 4 
+ Fe = FeS 0 4 -f- Cu. While the theoretical consumption of 
iron is 0.88 lb. for 1 lb. copper precipitated, the actual consump¬ 
tion was 1.5 times this amount. The solution was regenerated 
for further leaching by blowing hot air through the heated solu¬ 
tion, as shown by: ioFeS 0 4 -|— 5O = 3Fe 2 (S0 4 ) 3 -f- 2Fe 2 0 3 .S0 3 . 
Thus with a neutral solution, 40 percent of the iron content is 
precipitated as a basic salt ( 2 Fe 2 0 3 .S 0 3 ). To prevent the forma¬ 
tion of basic salt during the regeneration in the oxidizer, sul¬ 
phuric acid must be supplied in an amount as indicated by: 
2FeS0 4 + H 2 S 0 4 + O = Fe 2 (S 0 4 ) 3 -j- H 2 0 . Since acid must 
be replenished, the process becomes virtually one of leaching with 
acid. In treating mill tailings, 65 percent of the copper was 


TREATMENT OF COPPER ORE- 


215 


extracted in 3 hours (liquor boiled). In treating a io-ton lot 
of Cobre Grande ore containing 3 percent of copper, extraction 
was 96 percent. The cost per pound of copper was 6.4 cents. 
There was much trouble with the oxidizer in the regeneration 
of the spent liquor. (Greenawalt, p. 194; Mines and Methods, 
Sept., 1910, W. L. Austin). 

Thomas’s Experiments with Ferric Sulphate on Sulphide Ore. 

The results were: Double sulphides of copper, which occur 
in nature, require for complete transformation by ferric sulphate, 
long treatment with fine crushing. The commercial application 
does not pay. Free copper sulphides and oxides react easily with 
ferric sulphate in aqueous solution. Double sulphides of copper 
(such as chalcopyrite) must be roasted. Roast at low tempera¬ 
ture (sulphatizing), 450° C. to 480° C., to produce the maximum 
amount of copper sulphate. The copper may then be leached 
with but small amount of ferric sulphate. (Greenawalt, p. 201 ; 
Metallurgie, Jan. 15, Feb. 8, 22, 1904.) 

Methods of Extracting Copper at Rio Tinto, Spain. 

The ore is massive pyrite containing 3 percent of copper, mostly 
as cuprous sulphide. The method in use at the present time is 
weathering, or aid-oxidation. The reactions are: FeS 2 + 7O -f- 
H 2 0 = FeS 0 4 -f- H 2 S 0 4 . Ferrous sulphate readily oxidizes, 
forming ferric sulphate: 2FeS0 4 + H 2 S 0 4 + O = Fe 2 (S 0 4 ) 3 
-j- H 2 0 . Half the copper goes in solution in a few months, 
reducing an equivalent of ferric sulphate: Fe 2 (S 0 4 ) 3 + Cu 2 S = 
CuS 0 4 -J- 2FeS0 4 + CuS. The following reaction is slow, 
taking two years to extract 80 percent of the remaining half: 
Fe 2 (S 0 4 ) 3 + CuS + 3O + H 2 ,0 = CuS 0 4 + 2 FeS 0 4 + 
H 2 S 0 4 . The process is: Ore is piled in heaps containing 100,000 
tons, with flues of rough stone at the base connecting with 
chimneys. The oxidation goes on rapidly. The temperature 
rises, but is not allowed to go above 170° F. (77 0 C.) at chimneys, 
otherwise the heap might catch fire. The regulating of the tem¬ 
perature is accomplished by closing the tops of chimneys. Water 
is run on top of the heap, leaching the copper. The outflowing 
copper solution, before going to the precipitation tanks, is sent 
to the filter bed of fresh ore, thus reducing any contained ferric 
sulphate and insuring minimum consumption of iron : 7 Fe 2 (S 0 4 ) 3 

14 


1 


2l6 


ROBERT RHEA GOODRICH. 


+ FeS 2 + 8 H 2 0 = i 5 FeS 0 4 + 8 H a S 0 4 . After a year or 
two the copper contents is reduced to 0.3 percent, corresponding 
to 90 percent extraction. The ore is disposed of as “washed 
sulphur ore” (containing 49.5 percent S), which is used for the 
manufacture of sulphuric acid; 99.5 percent of the copper in 
the liquor is precipitated with a consumption of 1.4 lb. pig iron 
(92 percent Fe) per pound of copper recovered. (Greenawalt, 
p. 205; T. A. I. M. E., Voh XXXV, 1905, C. H. Jones.) 

Leaching Shannon Copper Ores. 

• 

Experiments were made by Francis S. Schimerka to find a 
practical leaching system for treating low-grade ore and tailings. 
The basic character of the ore makes sulphuric acid leaching 
impossible on account of the high acid consumption. Treatment 
with roasting gases gave successful results. The ore contained 
1.9 percent of copper, 0.55 percent of sulphur and a minute trace 
of gold and silver. The direct leaching with sulphuric acid gave 
an extraction of 81 percent, with acid consumption of 8.8 lb. 
per pound of copper extracted. The process employed was : The 
ore was subjected in heaps to roasting gases. At the base of the 
heap were constructed flues connecting with chimney for firing. 
At the bottom was placed a layer of 100 tons of pyrite containing 
50 percent of sulphur; then 1,000 tons of oxide ore, crushed to 
2 inch, was placed on top, and finally covered with a layer of 
fines a foot thick. The heap was sprinkled with waste solution 
from the scrap-iron precipitation tanks, producing ferric sulphate 
according to: 2 FeS 0 4 + S0 2 -f- 0 2 = Fe 2i (S 0 4 ) 3 . The roasted 
ore was leached with water in tanks with filter bottoms. The 
ferric sulphate formed by roasting dissolves copper which may 
not be water soluble. The outflowing copper solution, still con¬ 
taining some ferric sulphate, was run over tailings to reduce 
the last of the ferric sulphate before going to scrap-iron for 
precipitation of the copper. Extraction, 72 to 82 percent. The 
advantages of this process are: (1) Coarse crushing; (2) cheap 
installation; (3) no sulphuric acid used; (4) commercial use of 
ferric sulphate solution, produced in excess of requirements. 

Laboratory experiments by Francis S. Schimerka, on the treat¬ 
ment of ore and tailings, were conducted by roasting these mate¬ 
rials in a muffle furnace and treating, as follows: 


TREATMENT OE COPPER ORE. 


217 


1. Ore containing 2.37 percent Cu, 3.02 percent S', was roasted 
at 535 ° C. Leached with 5 percent H 2 S 0 4 solution. Extraction, 
86.4 percent. Consumption of acid, 3.19 lb. per pound of copper 
extracted—permissible in commercial practice. 

2. Ore containing 2.01 percent Cu, 2.58 percent S. Ferrous 
sulphate was added in such amount that iron (Fe) was 2.8 
percent the weight of ore. Leached with 10 percent H 2 S 0 4 
solution. Extraction, 82.5 percent. Consumption of acid, 1.2 lb. 
for 1 pound copper extracted. 

3. Tailings, product of concentration of sulphide ores (basic 
in character), containing 0.83 percent Cu, 0.88 percent S, were 
roasted with addition of pyrite in largest amount permissible for 
profitable working. When using sulphuric acid for the leach¬ 
ing the lowest consumption was 7.41 lb. acid per pound of copper 
extracted. Extraction 60 percent. 

4. Tailings, treated like sulphide ores. Ferrous sulphate 
added in such amount that iron (Fe) was 1.4 percent weight 
of ore. Roasted at 480° C. Ferric sulphate in roasted ore was 
1.01 percent. Water was added in amount equal to the weight 
of ore. Digested twelve hours with cold solution. Extraction 
71.7 percent. (E. and M. J., Vol. 96, No. 24, Dec. 13, 1913, 
p. 1107.) 

4. Chloride Processes. 

The chloride processes have been widely applied. Hydrochloric 
acid presents certain advantages over sulphuric acid in technical 
application. Hydrochloric acid is less apt to form basic salts, 
and therefore yields solutions that contain but little free acid 
and which constantly require less iron for the precipitation of 
copper. Hydrochloric acid is usually more expensive than sul¬ 
phuric acid. Ordinarily only oxidized ores are suitable to treat¬ 
ment by a chloride process. The copper may be dissolved by 
hydrochloric acid, or by a metal chloride. The reaction which 
occurs is: CuO -f- 2HCI = CuCl 2 -f- H 2 0 . (1.15 lb. HC 1 per 

pound Cu.) The precipitation of the copper is expressed by: 
CuCl 2 + Fe = FeCl 2 +. Cu. (0.88 lb. Fe per pound Cu.) 
Theoretically, the same amount of iron is required for the precipi¬ 
tation as for precipitation of copper from sulphate solutions. 
Practically, sulphate solutions consume more iron than chloride 
solutions. (Greenawalt, p. 216.) 


2l8 


ROBERT RHEA GOODRICH. 


In Stadtberg, Westphalia, hydrochloric acid was formerly used 
to extract copper from ores containing i to 2 percent of copper. 
The process replaced a sulphuric acid process previously em¬ 
ployed, but was abandoned when carbonates of the ore changed 
to sulphides in depth. The ore was leached in rectangular wooden 
tanks containing 90 tons. The fresh ore was treated with par¬ 
tially saturated solution until saturation took place. Ore nearly 
copper-free was treated with fresh 12.5 0 B. acid, and lastly with 
water wash. Acid consumption, slightly above the theoretical. 
(Schnabel, ‘‘Handbook of Metallurgy/’ Vol. I, p. 200; Greena- 
walt, p. 217.) 

Ferric chloride, like ferric sulphate, dissolves copper from ores 
containing the metal as oxide, carbonate and sulphide. (Cu 2 S is 
the only sulphide easily soluble.) The reactions are: Cu 2 S + 
4FeCl 3 = 2 CuC 1 2 + 4 FeCl 2 +, S, and 3C11O + 2FeCl 3 = 3 CuC 1 2 
+ Fe 2 O s . The copper may be precipitated by iron: CuCl 2 -j- Fe 
= FeCl 2 + Cu. Regeneration of the ferric solution may be 
accomplished by means of air, causing one-third of the iron to 
precipitate: 6FeCl 2 -j- 3O = 4FeCl 3 -f- Fe 2 0 3 . It may be 
regenerated by chlorine, produced chemically or electrolytically: 
FeCl 2 —(— Cl = FeCl 3 . (Greenawalt, p. 218.) 

Doetsch Process. 

This process was formerly applied to raw and to roasted ores 
at Rio Tinto, Spain. 

1. Raw ores. The ore contained 2.7 percent of copper. The 
pyrite was unaffected by leaching solutions. The process is: 
Crush to ]/2 in. Mix with 0.5 percent by weight of sodium 
chloride and 0.5 percent of ferrous sulphate. Build in large 
heaps. Run on solution of ferric chloride. Precipitate copper 
on pig iron. Regenerate solution in scrubbing towers by chlorine 
generated by roasting, according to: 2FeS0 4 + 4 NaCl -j- 3O = 
Fe 2 0 3 + 2Na 2 S0 4 -j- 4CI. Extraction, 50 percent in 4 months, 
80 percent in 2 years. Consumption of pig iron, 1.3 lb. per pound 
of copper produced. Cumenge estimates the cost as 3.3 cents 
per pound of copper. 

2. Roasted ores. The ore of nut size is made into heaps of 
800 tons, with 14 tons salt. Rough flues 20 inches square, used 
for firing, are constructed at the bottom and connected with 


TREATMENT OF COPPER ORE. 


219 


chimneys. The reaction is shown by: 2FeS0 4 + 4NaCl + 3O 
= Fe 2 0 3 + 2Na 2 S0 4 + 4CI. The liberated chlorine produces 
ferric chloride and cupric chloride. Some of the roasted ore has 
been mixed at times with the unroasted ore for the extraction of 
the copper by the ferric chloride solution. (Greenawalt, p. 219; 
Annales des Mines, Vol. XCVI; Notes Sur le Rio Tinto, M. E, 
Cumenge.) 

The Froelich Process. 

The ore is subjected, in the absence of air, to temperature 
betweeri 150° C. and 8oo° C. Chlorine gas is introduced. After 
chlorinating, in order to regain the chlorine combined with the 
iron, heat to above 300° C., introducing a certain amount of 
air, when the ferric chloride is evaporated, and in the presence 
of air is decomposed into ferric oxide and chlorine. The cupric 
chloride is not changed by this treatment. Leach with water. 
Precipitate copper by iron. Oxidize ferrous chloride to ferric 
chloride in rotating drum by a blast of air. Drive off water of 
crystallization of ferric chloride. Heat to about 300° C. with a 
certain amount of air, thus regenerating chlorine for the process. 
(Greenawalt, p. 223; U. S. Patent No. 846,657, Mar. 12, 1907.) 

Ferrous Chloride Process. 

Some years ago* Hunt and Douglas based a process on the 
action of ferrous chloride on copper oxide and copper carbonate. 
Hunt says that chrysocolla (CuSiOg -|- 2H 2 0) is likewise com¬ 
pletely decomposed by a hot solution of ferrous chloride contain¬ 
ing sodium chloride. The reaction given is: 3CuO + 2FeCl 2 — 
CuCb + 2C11CI + Fe 2 0 3 . The precipitation by metallic iron 
is said to be according to: CuCl 2 + 2CuCl + 2Fe = 2FeCl 2 
-F 3C1U Cuprous chloride, which is insoluble in water, is soluble 
in solutions containing an excess of metal chlorides. Silver in 
ore is converted to silver chloride by cupric chloride, and is 
dissolved in solutions containing an excess of metallic chlorides. 
Theoretically, 0.59 lb. iron precipitates 1 lb. copper. An objec¬ 
tion is that basic salts of iron clog the filter. Heat is not neces¬ 
sary, but it hastens the reaction. The silver will be precipitated 
with the copper. If it is desired to precipitate the silver sepa¬ 
rately, reduce all cupric chloride to the cuprous state by means 


220 


ROBERT RHEA GOODRICH. 


of sulphur dioxide, then precipitate the silver on copper, and 
subsequently copper on iron. 

The process, as formerly applied at Ore Knob, Ashe Co., N. C., 
is: Crush the ore to 40 mesh and roast to sulphate and oxide. 
Leach with hot solution of ferrous sulphate and sodium chloride 
(22 0 B.). Precipitate copper on iron, regenerating solvent: 
0.7 lb. iron precipitates 1 lb. of copper. Cost of the copper, 8 
cents per pound. (Greenawalt, pp. 225-6; T. A. I. M. E., Vol. X, 
p. 12; T. A. I. M. E., Vol. II, p. 394, E. E. Olcott.) 

Hunt and Douglas Process. 

This process is based on the following combination reaction: 
2 CuS 0 4 + zNaCl + S 0 2 + zH 2 0 = 2C11CI .+. Na 2 S 0 4 + 
2 H 2 S 0 4 . The process is: Roast, if the ore is a sulphide, and 
extract the copper from the oxidized ore with dilute sulphuric 
acid regenerated in the process. Convert the cupric sulphate 
into cupric chloride by addition of a soluble chloride. Convert 
soluble cupric chloride into insoluble cuprous chloride by sulphur 
dioxide, with simultaneous regeneration of sulphuric acid. Con¬ 
vert the precipitated cuprous chloride into cuprous oxide, or 
metallic copper, by addition of milk of lime or replacement with 
iron. If the roasted ore contains silver, the sulphate of copper, 
which in well-roasted ore should be one-third of the content, is 
first leached out with water, care being taken to add just sufficient 
soluble chloride to render insoluble any silver present. From 
the clear solution thus obtained, after adding the requisite amount 
of sodium chloride to chloridize the copper, precipitate the copper 
by sulphur dioxide. The resulting liquor, freed from sulphur 
dioxide, is used to dissolve the oxide of copper in the ore. From 
the residues the silver may be extracted by brine, after which 
the gold may be recovered by chlorination. Chloride of silver 
is soluble to some extent in a solution of cupric chloride, and 
is then in part carried down with the cuprous chloride. (Greena¬ 
walt, p. 228; T. A. I. M. E., Vol. X and XVI.) 

Hunt and Douglas Process at Argentine, Kansas. 

At the Kansas City Smelting and Refining Company’s works 
the material treated was a lead-copper matte. In applying the 
process, using sodium chloride as chloridizer, sodium sulphate 


TREATMENT OF COPPER ORE. 


22 I 


accumulated in the solution, which crystallized out, causing 
trouble. This trouble was corrected by using three parts of 
calcium chloride and one part of sodium chloride as chloridizers. 
Calcium sulphate formed, and, being insoluble, was precipitated. 
The necessary calcium chloride was produced in the process, 
in the conversion of cuprous chloride to cuprous oxide by milk 
of lime. The residue from the leaching went to the lead smelting 
department. (Greenawalt, p. 231 ; Mineral Industry, Vol. XVII, 
1908, p. 296.) 

Modification of the Hunt and Douglas Process. 

To simplify the operation, Hofmann worked out and success¬ 
fully introduced the following modifications: Roast ore or matte 
as usual. Treat roasted ore or matte in agitating tanks with 
dilute sulphuric acid produced in the process: H 2 S 0 4 -f- CuO = 
CuS 0 4 H a O. Chloridize with hydrochloric acid produced in 
the process, regenerating sulphuric acid: CuS 0 4 + 2HCI — 
H 2 S 0 4 -f- CuCl 2 . Convert the soluble cupric chloride into 
insoluble cuprous chloride by cement copper in an agitating tank. 
Heat by a steam jet, forming cuprous chloride: CuCl 2 + Cu — 
2CuCl. Treat the cuprous chloride in revolving barrels with a 
small amount of water, scrap iron and salt (sodium chloride). 
The salt helps to start the reaction by dissolving some cuprous 
chloride: 2CuCl -f- Fe = FeCl 2 + 2Cu. Evaporate the ferrous 
chloride solution in an iron pan. Charge the solid ferrous chlo¬ 
ride into retorts which are provided with water (for steam) and 
air, hydrochloric acid being regenerated for use in the process: 
2 FeCl z -f- O + 2H0O = Fe 2 O s + 4HCI. This modified process 
was used for some time until Hofmann received instructions to 
change the plant over to the more profitable manufacture of blue 
vitriol. (Greenawalt, p. 241; Mineral Industry, Vol. XVII, 
1908, p. 296, Ottokar Hofmann.) 

The Bradley Process. 

The sulphide ore is carefully roasted at 450° C. to 550° C. 
in a roasting furnace known as the “amphidizer.” The roasted 
ore is treated with calcium chloride solution in a reaction drum 
at ioo° C., converting copper sulphate into cupric chloride, and 
any ferric sulphate produced by roasting is changed to ferric 


222 


ROBERT RHEA GOODRICH. 


chloride. The resulting calcium sulphate is insoluble. The ferric 
chloride which forms dissolves copper oxide, copper sulphide 
and metallic copper which may have remained unaffected by 
roasting. The gold and silver in the ore are brought into solu¬ 
tion by adding free chlorine. The chlorides of silver and gold, 
soluble in calcium chloride solution, may be precipitated with the 
copper. After leaving the reaction drum the pulp is filtered. The 
solution is subjected to further oxidizing operations so as to be 
sure the metals are all combined at their highest valency. The 
solution is then in condition for treatment for the separation of 
the dissolved metals. (Greenawalt, p. 243; E. and M. J., Jan. 6, 
1912; U. S. Patent, No. 1,011,562, Dec. 12, 1911.) 

Longmaid-Henderson Process. 

This process consists in roasting pyrite cinder from sulphuric 
acid works with sodium chloride, leaching out the cupric chloride 
formed and precipitating the copper by iron. Longmaid obtained 
patents, Oct., 1842, and Jan., 1844, relating to the treatment of 
pyrite cinders by roasting with sodium chloride. He may be 
regarded as the pioneer in the field of the wet extraction of 
copper. Gossage, in 1850, first used sponge iron to precipitate 
copper. Henderson improved the process in 1865, adding absorp¬ 
tion towers, so that the gases from chloridizing roasting yielded 
weak acid, which was used for leaching copper ores. The steps 
in the process are: Mix the roasted ore (pyrite cinder) with 
sodium chloride, and grind mixture. Chloridizing roasting. 
Leach the roasted ore. Precipitate the silver from the argen¬ 
tiferous liquors. Precipitate the copper from the desilverized 
liquor. Prepare the leached ore for the iron works. Analysis 
of average pyrite cinder, 4 to 5 percent copper; sulphur, equal to 
or greater than copper. Mechanical furnaces use less salt than 
hand roasters, average 7.5 percent. 

In the German works at Oker, average results showed: 75 
percent of the copper soluble in water, 20 percent of the copper 
soluble in dilute hydrochloric acid, 5 percent of the copper 
insoluble. The leaching of the roasted chloridized ore was done 
in wooden tanks with filter bottoms of perforated boards painted 
with coal tar. The process was: Weak liquor from previous 
tank was applied, going off as strong liquor to precipitation tanks. 


TREATMENT OF COPPER ORE. 


223 


i hen two hot washes were applied, producing weak liquor to be 
used again on another tank. Next, dilute tower acid was put on 
six times; a satisfactory extraction resulted. The acid liquors, 
which produce somewhat impure copper, may be precipitated 
separately. Copper precipitated from neutral solution is pure. 
Copper precipitated from acid solution is impure. Most cup¬ 
riferous pyrite contains some silver and small quantities of gold. 
Silver chloride soluble in other chlorides, and gold chloride soluble 
in water, are extracted with the copper. Formerly there were 
various schemes proposed for the separate recovery of gold and 
silver from copper. Now usually there is no attempt made to 
recover gold and silver separately from copper, because of the 
cheap electrolytic method of refining copper. The residue, “pur¬ 
ple ore," was briquetted and sent to the blast furnace. (Greena- 
walt, p. 246; Lunge’s Treatise on the Manufacture of Sulphuric 
Acid and Alkali, Vol. I, 1891.) 

The Longmaid-Henderson process as carried out at the works 
of the Pennsylvania Salt Manufacturing Company at Natrona, 
Pa., where 200 tons of pyrite cinder is treated per day, is: Grind 
to 20 mesh in Chilean mill with 10 percent of salt. Add raw ore 
so that sulphur equals ifT times weight of copper. Charge 9,600 
lb. mixture into the muffle roasting furnace, kept 8 hrs. at dull red 
heat, 800 0 C. Send roasted material to cooling floor, then to 
leaching vats. It is not desirable tO' precipitate silver by Claudet’s 
iodide process, because it leaves 5 oz. silver to the ton of copper, 
and electrolytic refineries and blue vitriol works will pay for 95 
percent of the silver and all the gold and copper. Cost to treat 
one ton of pyritic cinder, $1.87. (Greenawalt, p. 262; Mineral 
Industry, Vol. VIII, IX; Joel G. Clemer.) 

The cost of producing copper by the Longmaid-Henderson 
process in a modern plant, using mechanical roasters: Eastern 
U. S., treating per day 30 tons pyrite cinder containing 2.27 per¬ 
cent Cu, 2.28 percent S, was 5.2 cents per pound of copper. 
Western U. S., under similar conditions, it was 6.5 cents per 
pound of copper. (Greenawalt, p. 266.) 

Copper Precipitated by Chemical Reagents. 

The copper precipitants used are: 1. Iron : Cast iron, wrought 
iron, sponge iron. 2. Hydrogen sulphide. 3. Lime. 


224 


ROBERT RHEA GOODRICH. 


1. Iron precipitates copper according- to: CuS 0 4 + Fe = 
Cu + FeS 0 4 ; CuCl 2 + Fe = Cu + FeCl 2 . (Theoretically, 
from solution of cupric salts, 0.88 lb. of Fe precipitates i lb. 
of Cu.) 2CuCl -j- Fe — 2Cu + FeCl 2 . (From solutions of 
cuprous chloride, 0.44 lb. Fe precipitates 1 lb. Cu.) Iron may 
be consumed by ferric salts and by free acid: Fe 2 (S 0 4 ) 3 -f- Fe 
= 3FeS0 4 ; 2FeCl 3 + Fe = 3FeCl 2 ; H 2 S 0 4 + Fe = FeSQ 4 
-j- H 2 . Thus for minimum iron consumption the solution going 
to precipitation (“cementation”) tanks should be free from ferric 
salts and low in acid. 

At Stadtberg, Westphalia, where hydrochloric acid was used 
for leaching, 1.27 lb. iron precipitated 1 lb. copper. In the Flunt 
and Douglas process, 0.5 lb. to 0.7 lb. iron precipitated 1 lb. copper 
from cuprous chloride solutions. At Gumeskevsky Mine, Russia, 
where cast iron was used both in plates and granulated, it was 
found that tanks charged with 12 tons of granulated cast iron, 
placed in layers 4 inches thick on four inclined false bottoms, 
gave as good precipitation as tanks charged with 115 tons of 
plates. Wrought iron scrap is much used. 

It is often advantageous to make the iron on the spot. Sponge 
iron may be made by heating ferric oxide in a reducing atmos¬ 
phere and cooling in a reducing atmosphere. In a measure it 
solves the problem of iron supply. It has long been known that 
iron sponge can be thus produced, but the idea has not met with 
much encouragement. 

2. Hydrogen sulphide precipitates copper as cupric sulphide: 
CuS 0 4 + H 2 S = CuS + FI 2 S0 4 ; CuCl 2 + H 2 S = CuS + 
2FICI. Acid regenerated by the precipitation is returned to the 
ore to dissolve more copper. This does not regenerate all the 
acid applied to the ore; some acid must be supplied to the 
process. 

3. Lime (calcium hydrate or milk of lime) is not suitable for 
precipitating copper from sulphate solutions on account of con¬ 
taminating the precipitated copper with calcium sulphate. It was 
for a long time used with the Hunt and Douglas process, and 
was found to be a successful precipitant because the resulting 
calcium chloride is very soluble: 2CuCl + Ca(OH) 2 = Cu 2 0 
+ CaCl 2 .+ H 2 0 . (Greenawalt, p. 270 to 282; Lunge, “Sul¬ 
phuric Acid and Alkali Manufacture,” p. 815.) 


treatment of copper ore. 


225 


II. Electrolytic Methods: 

1 he successful electrolytic refining of blister copper has turned 
the attention of metallurgists to the application of similar opera¬ 
tions in the extraction of copper direct from its ores. There are 
difficulties in the electrolytic extraction of copper from its ores 
that are not met with in the electrolytic refining of copper, yet 
these difficulties do not appear insurmountable. 

In electrolytic refining a soluble anode is used. Theoretically, 
there is no consumption of energy, and no acid consumed or 
generated. With the soluble anode there is no serious difficulty. 
When copper is dissolved from the ore, conditions are different. 
The insoluble anode is the first serious difficulty. It is hard to 
find a good conducting substance which is not attacked during 
electrolysis. Graphitized carbon has given satisfaction with 
chloride electrolytes, but not with sulphate electrolytes. For 
sulphate solutions no satisfactory anode has been found; on 
the whole, peroxidized lead is the best material for this purpose. 

The cathodes are usually thin sheets of pure copper. In 
depositing copper from impure solutions derived from the leach¬ 
ing of ores it is found difficult to get a reguline deposit, unless 
considerable care is taken with purifying the electrolyte and the 
regulation of current density. Irregular deposition and sprout¬ 
ing may require the removal of the cathode before acquiring 
the desired thickness. With reasonably pure electrolyte and low 
current density, difficulty will not occur, especially if the electro¬ 
lyte is agitated or the cathode oscillated. The current density and 
the nature of the electrolyte have much to do 1 with the purity 
and the quality of the deposit. 

Most electrolytic processes are based on the regeneration of 
the solvent during electrolysis. This requires that the anolyte 
and catholyte be kept separate. A diaphragm which will allow 
the current to pass, but prevent solutions from mixing, becomes 
necessary. Few materials. for diaphragms fulfill all of the 
requirements. They must be permeable to ions, prevent diffu¬ 
sion of electrolyte, and carry no current by metallic conduc¬ 
tion. Asbestos is the best material, as it is not easily attacked 
by acid or alkaline solutions. It is used as cloth, paper or 
mill board. 

The current density affects the efficiency of the operation and 


226 


ROBERT RHEA GOODRICH. 


the nature of the copper deposited. It is not possible to use as 
high current density as in electrolytic refining, because of the 
impure and leaner solution. To get reguline deposit, operation 
at low current density is necessary. High current density causes 
impoverishment of ions at the electrodes, which may be prevented 
by stirring the electrolyte or by a moving electrode. Energy 
efficiency becomes less as the current density increases. 

Wilde was one of the first to deposit copper on a revolving 
cathode. He secured an even distribution of the copper. The 
current density was 20 amperes per sq. ft. (2.2 per sq. dm.). 
Elmore used horizontal mandrels. The current density was 30 
amperes per sq. ft. (3.3 per sq. dm.). Copwer-Coles used a 
cylindrical cathode, revolving at a speed of 1500 to 2000 lin. ft. 
(450 to 600 m.) per minute, with a current density of 200 
amperes per sq. ft. (22 per sq. dm.). (U. S. Patent 895,163, 

Aug. 4, 1908, Cowper-Coles; Greenawalt, p. 283 to; 290.) 

Theoretical Data. 

1 ampere-hour deposits from cupric solution, 1.1858 gr. copper. 

1 ampere-hour deposits from cuprous solution, 2.3717 gr. 
copper. 

12,000 ampere-hours deposits from cupric solution, approx¬ 
imately 32 lb. copper (theoretically 31.15 lb.). 

12,000 ampere-hours deposits from cuprous solution, approx¬ 
imately 64 lb. copper (theoretically 62.30 lb.). 

The theoretical voltage required for electrolyzing with insoluble 
anodes is: For cupric sulphate, 1.22 volts; cupric chloride, 1.35 
volts; cuprous chloride, 1.53 volts. 

The output in depositing copper from solutions with insoluble 
anodes is: 

A 

2.1429 lb. per kw. hour. 
1.9364 lb. per kw. hour. 
3.4174 lb. per kw. hour. 


Cupric sulphate, - — 


12 X 1.22 


Cupric chloride, -^ 


12 X 1.35 


Cuprous chloride, —— - 

12 X 1.53 





TREATMENT OF COPPER ORF. 


227 


Illustration: There were used in precipitating copper from 
cuprous chloride solution 400 amperes per 12 hours (4800 ampere- 
hours) at 1.8 volts. Copper deposited, 18.2 pounds. Theoreti¬ 


cally there should be deposited - 4 8qq x 64 = 25.6 lb. 

12,000 

• X 3 2 

Current efficiency, —to = 71.2% (practice, about 90%). 


There was deposited, 


25.6 

18.2 


4.8 X 1.8 


2.11 lb. of copper per kw. hour. 


Energy efficiency, M!_ = 61.8% (practice, about 50%). 

3.417 - 

For electrolytic refineries, Addicks gives a rough summary of 
the relative value of the resistances in practice: Metallic resistance 
15 percent, electrolytic resistance 60 percent, contacts 20 percent, 
counter E. M. F. 5 percent. The counter E. M. F. in copper 
refining, due to greater concentration at the anode than at the 
cathode, is 0.02 volt. Copper refining employs 0.2 to 0.4 volt 
between the electrodes. Copper depositing with insoluble anodes 
employs 1.5 to 3.0 volts, depending on the current density and 
concentration of the electrolyte. (Greenawalt, p. 295; The 
Journal of the Franklyn Institute, Dec., 1905, Addicks.) 

The solvents usually employed have sulphuric or hydrochloric 
acid as the basis. The cycle consists of: Solution, precipitation, 
regeneration. In a cyclic process the solution is likely to become 
charged with impurities and reduce the efficiency of the deposition. 

The injurious effects of impure electrolyte are: 

1. Undesirable metals may be deposited with copper, when 
the solution becomes impoverished. 

2. Useless energy is expended in, reduction and oxidation, 
deposition and immediate dissolving. 

The metals whose compounds have the lowest heats of forma¬ 
tion are first deposited. Gold, silver and copper come first, and 
are deposited in the order given. If the current density and, 
consequently, voltage exceed a certain amount, several metals may 
be deposited together. The more neutral the electrolyte, the more 
likely the more electro-positive metals (such as iron, nickel and 
zinc) are to be deposited. High current density may deposit 
copper and zinc together from slightly acid electrolytes. If 






228 


ROBERT RHEA GOODRICH. 


much acid is present, hydrogen will be liberated, causing low 
efficiency. High current density may deposit copper and a more 
electro-positive metal not readily redissolved, producing impure 
copper. The practical factors which determine the kinds of 
ions deposited at the cathode are: Heats of formation of the 
constituents of the electrolyte, concentration of anolyte and catho- 
lyte, current density at cathode, temperature of electrolyte. 

If arsenic and antimony were originally in the ore, they should 
have been largely eliminated by roasting. If they should accumu¬ 
late in the solution, they may be precipitated by hydrogen sul¬ 
phide, which, however, is expensive. Iron is most likely to be 
in the lixivium. In sulphate solutions the iron may accumulate 
to saturation, unless the solution is purified at intervals. With 
chloride solutions, more or less ferric oxide is precipitated by 
reaction with the ore. The influence of iron in the electrolyte 
may be summed up as follows: 

1. Becomes oxidized to ferric salt at the anode. 

2. Is carried to cathode by diffusion and circulation. 

3. Dissolves precipitated copper, becoming reduced to the 

ferrous condition. 

4. Is carried again to the anode by diffusion and circulation, 

again becoming ferric iron. 

This cycle continues indefinitely, greatly reducing the efficiency 
of the electric current. To economically precipitate copper from 
an electrolyte containing much iron, either a diaphragm must be 
used to prevent diffusion, or a reducing agent (acting as a 
depolarizer) must be introduced. 

There are two alternatives relative to impure electrolyte : 

1. Purification of electrolyte. 

2. Wasting a portion of the solvent, and replacing with fresh 

solution. 

Ulke states that the best method is to electrolyze in special 
vats with lead anodes and to use a current density sufficiently 
strong to deposit arsenic and antimony, but not strong enough to 
deposit iron. This is repeated till the bath contains so much 
iron that it is necessary to remove it by crystallizing out the 
ferrous sulphate. (Greenawalt, p. 300; “Modern Electrolytic 
Copper Refining,” Titus Ulke.) 


TREATMENT OE COPPER ORE- 


229 


Ottaker Hofmann gives the following method for purifying 
copper sulphate solutions containing as impurities salts of iron, 
arsenic, antimony, bismuth, etc.: The crude copper sulphate 
solution is forced into towers lined with lead and provided with 
a lead steam coil for heating. A lead pipe connected with an air 
compressor enters through the funnel-shaped bottom. When 
the solution is hot, roasted matte is added and air forced in, 
causing precipitation of the iron according to : 2FeS0 4 + O + 
2C11O = Fe 2 O s + 2 CuS 0 4 . To observe and regulate the prog¬ 
ress of the operation the solution is tested for iron from time 
to time. When the solution is free from iron it will contain no 
trace of other impurities. (Mineral Industry, Vol. VIII, p. 192, 
Ottaker Hofmann.) 

At the Kalakent Copper Works, Russia, impure electrolyte 
was run over dead roasted matte, then over a heap of low-grade 
copper ore. The neutral solution was run into a lead-lined vat, 
where it was diluted to 1 12 0 B. and heated to ; 50° C. Scrap 
anode copper was hung in the vat to neutralize any acid found 
in the operation. Compressed air was forced in till the concen¬ 
tration was 15 0 Be., the scrap copper being quickly dissolved. 
The liquid was drawn off and clarified. The excess of purified 
electrolyte which gradually accumulated was withdrawn from 
the system and worked up into' bluestone (CuS 0 4 . 5 H 2 0 ). The 
copper produced was 99.9 percent pure. (“Modern Electrolytic 
Refining,” p. 145, Titus Ulke.) 

Copper sulphate may be crystallized out of impure solutions, 
redissolved in water and electrolyzed to deposit the copper, 
thereby regenerating acid which is applied to leaching the ore. 
The impurities are thus eliminated. 

Greenawalt proposes purifying chloride solutions by electro¬ 
lyzing sodium chloride, producing free chlorine and caustic soda: 
NaCl -j- H 2 0 + electric current = NaOH + Cl + H. Caustic 
soda is applied to a portion of the withdrawn electrolyte, causing 
precipitation of the bases according to: RC 1 2 + 2NaOH = 
2NaCl + R(OH) 2 . (R represents the base metals.) The 
chlorine is converted into hydrochloric acid: 2CI + S 0 2 + 2 H 2 0 
— 2HCI + H 2 S 0 4 . Thus the impurities are eliminated and the 
hydrochloric and sulphuric acid solvents are regenerated. 
(Greenawalt, p. 303, p. 352.) 


230 


ROBERT RHEA GOODRICH. 


If impurities in the electrolyte do not materially interfere with 
the efficiency of the process, it is better to work with impure 
solution even though impure copper be deposited. The efficiency 
of the process is more important than the relative purity of the 
copper. 

Cobly, in 1878, described the use of sulphur dioxide for the 
depolarization of insoluble anodes in the deposition of copper. 
Its operation in recent years is common and well understood. 
Other depolarizers have been suggested. (Greenawalt, pp. 283 
to 309.) 

The following theoretical voltages were calculated by intro¬ 
ducing Richards’ thermochemical data in reactions, which we 
believe represent the operation of the electrolytic cells: 

I. Deposition of copper from sulphate solution: 

(a) Without using a depolarizer, CuS 0 4 + H 2 0 + elec¬ 
tric current = Cu + H 2 S 0 4 + O — 56,300 calories. 


Theoretical voltage, 7— -———. = 1.22 volts. 

(96540 X 0.24 X 2) 

(b) Using S 0 2 gas as a depolarizer, CuS 0 4 -f- S 0 2 + 
2H0.O + electric current = Cu + 2H 2 S0 4 + 7,300 calories. 

Theoretical voltage, j— --——r = —0.15 volt. 

- (96,540 X 0.24 X 2) J 

(c) Siemens and Halske process, using FeS 0 4 as depolar¬ 
izer, CuS 0 4 + 2FeS0 4 + electric current = Cu + Fe 2 (S 0 4 ) 3 
—16,800 calories. 

Theoretical voltage, 7- 7 _ l6 ’ So ° — 0.36 volt. 

(96,540 X 0.24 X 2) 0 

II. Deposition of copper from cupric chloride solution: 

(a) Without using a depolarizer, CuCl 2 -j- electric current 
= Cu 2CI —62,500 calories. 


Theoretical voltage. 


62,500 


= !-35 volts. 


(96,540 X 0.24 x 2). 

(b) Body’s process, using FeCl 2 as a depolarizer: 2FeCL 
CuCl 2 —)— electric current = = Cu —|— 2FeCl 3 —7,000 calories. 


Theoretical voltage, /-- 

(96,540 X 0.24 X 2) 


= o.iq volt. 









TREATMENT OF COPPER ORE. 


231 


III. Deposition of copper from cuprous chloride solution: 

(a) Without using a depolarizer, CuCl + electric current 
: Cu -j- Cl —35,400 calories. 


Theoretical voltage, 


35,4oo 


1 .53 volts. 


(96,540 X 0.24) ' 

(b) Hoepfner's process, using CuCl as a depolarizer, 
2CuCl -j- electric current = Cu + CuCl 2 —19,400 calories. 


Theoretical voltage, 


19,400 


= 0.84 volt. 


(96,540 X 0.24) 

In the above equations, where the minus sign (—) precedes 
the numerical value of the calories, this sign means that the 
given number of calories is absorbed in the reaction and that 
their equivalent in electrical energy must be supplied from an 
external source, viz., the dynamo. Thus the calculated voltage 
is the necessary theoretical impressed electromotive force that 
must be supplied in order that the reaction may take place. 
When a depolarizer is used the required impressed volts are lower 
than when a depolarizer is not used. Compare equation I (c), 
0.36 volt, with equation I (a), 1.22 volts. There is thus a 
saving of energy and also of expense in precipitating copper 
when using a depolarizer. This reducing of voltage due to using 
a depolarizer is realized more efficiently with low current density. 
The gas is then liberated at the anode at such a moderate rate 
that it will have a chance to react with the depolarizer, and not 
escape as such without reacting. In proportion as this reaction 
is more complete, so will the theoretical voltage be more nearly 
attained. 

Referring to an experiment of K. Reinartz, recorded by Austin, 
Reinartz secured an anode efficiency of 65 percent when using 
sulphur dioxide as a depolarizer. Thus 65 percent of the oxygen 
liberated at the anode did react with the sulphur dioxide 
depolarizer. (“Metallurgie,” 1908, pp. 202 to 205.) 

Equation I (b) works out with a ( + ) sign for the calories 
and gives (—0.15) volt. This would indicate that with such 
a current density as to secure anode efficiency of 100 percent, 
if this could be realized, the electrolytic cell would no longer 
require an impressed voltage, but would become a primary cell. 
Of course, a dynamo would have to supply voltage to realize 


15 




232 


ROBERT RHEA GOODRICH. 


a current of any magnitude, but the cost of operating, were this 
high anode efficiency secured, would be as low as that of elec¬ 
trolytic refining. (Mines and Methods, Aug., 19 11 , P- 2 & 2 > W. L. 
Austin, Reinartz Experiment.) 

1. Electrolytic Sulphate Processes. 

There are two general classes of processes based on the sol¬ 
vents sulphuric acid and ferric sulphate. Neither of these solvents 
can be employed to the exclusion of the other. On certain 
copper sulphide ores, ferric sulphate solution has given good 
results without roasting. The necessity of using a diaphragm, 
or the introducing of a reducing gas such as sulphur dioxide 
when iron salts are present, has been mentioned. 

Sulphuric Acid Process. 

Only oxidized or roasted ores are effectively treated by dilute 
sulphuric acid. See equation I (a), where the theoretical voltage 
= 1.22 volts. The minimum voltage required is 1.22 volts. In 
practice 1.5 to 3 volts are employed. There is regenerated 
sulphuric acid in an amount equivalent to the copper deposited. 
This acid will only partially supply the acid required for leaching 
ore, because a certain amount of acid is consumed by action on 
the gangue. Yet electrolytic precipitation is an advance over 
precipitation by metallic iron, in which case no acid is regener¬ 
ated. Theoretically, 2.1429 lb. copper are deposited per kw. hour. 
This energy efficiency of 100 percent cannot be attained, because 
it would require the ohmic resistance of the entire circuit to be 
nil. Fifty percent energy efficiency may be expected. When 
using sulphur dioxide as depolarizer (see equation I (b)), where 
theoretical voltage equals —0.15 volt, no power will be required 
theoretically to deposit the copper, and the process will be placed 
on a par with electrolytic refining with a soluble anode. The 
practical power consumption depends upon the ohmic resistance 
of the circuit and the current density employed. Tossizza ascer¬ 
tained by experiment that by using sulphur dioxide as depolar¬ 
izer the necessary impressed voltage is diminished to 0.2 volt. 
Electrolytic refining uses 0.2 to 0.4 volt. (Greenawalt, p. 309; 
U. S. Patent 710,346, Sept. 30, 1902, Tossizza.) 


TREATMENT OF COPPER ORE). 


233 


Treatment of Ore by Braden Copper Co. } Chile. 

The plant consists of a modern concentrator, a smelter and 
a leaching plant capable of treating one-fifth of their output of 
concentrates. They estimated that on account of: local conditions 
(cheap water power, etc.) the leaching of concentrates would be 
cheaper than smelting. The concentrates contain 16 percent Cu, 
19 percent Fe and 22 percent S. The process is: Roasting con¬ 
centrates; manufacturing sulphuric acid for leaching; precipi¬ 
tating the copper electrolytically. (Greenawalt, p. 331; E. and 
M. J., Dec. 30, 1911, Pope Yeatman.) 

The Chile Exploration Company, Chuquicamata. 

The copper in this ore exists as brochantite (CuS0 4 .3Cu(0H) 2 ), 
which is insoluble in water, but soluble in dilute sulphuric acid. 
There is no arsenic, antimony or silver in the ore. The first 
25 ft. of the ore body contains 0.02 percent sodium chloride. 
A difficulty first met was that during electrolysis obnoxious 
chlorine gas was given off. The process developed by A. E. 
Cappelen Smith is: Crush to 34 to 34 inch. Leach in large open 
vats with dilute sulphuric acid. (Percolation, 16 ft. high.) Drain 
and wash with water. Run copper liquor, prior to' electrolysis, 
over shot copper in order to> eliminate the chlorine. The cuprous 
chloride thus precipitated is smelted at a low temperature with 
coke and lime. Calcium chloride is slagged. There is no loss 
of copper by volatilization during smelting of the chloride. 
Deposit the copper electrolytically, using insoluble magnetic oxide 
anodes (Fe 3 0 4 ). There is a gain of 5 kilos of sulphuric acid per 
ton of ore treated. Run 8 to 10 percent of the liquor to waste. 
Extraction, 90 percent. This is a leaching plant which is being 
operated under favorable conditions. The power is transmitted 
from the seacoast, 41 miles distant. (E. and M. J., Oct. 4, 1913, 
Vol. 96, No. 14, p. 651 ; Met. and Chem. Eng., May, 1914, XII, 
p. 291.) 

The Butte-Duluth Mining Company, Butte, Montana. 

The ore is a decomposed granite, containing 2 percent of copper 
as malachite, azurite, chrysocolla and cuprite. The mining is by 
open cut method. The process is: Crush to 34 inch. Send to 
leaching tanks. Leach with 10 percent sulphuric acid solution, 


234 


ROBERT RHEA GOODRICH. 


put on for 24 hours. Pass liquor through the temperature cells, 
where it is heated to 6o° C. Send heated solution to electrolytic 
cells. Send electrolyzed lean solution to sump, where it is stand¬ 
ardized to 10 percent of H 2 S 0 4 . Return the replenished solution 
to leaching tanks. In washing, make first water wash small in 
quantity, and add to mill solution. Waste an equivalent amount 
of the mill solution, which, prior to wasting, is run over ore 
till neutral, and over scrap iron to' precipitate the copper. With 
the six new electrolytic cells, output is 100 tons finished copper 
(96.96 percent pure) per month. Acid consumption, 3.5 lb. per 
pound of copper. One pound copper per kw. hour. Cost, 13.7 
cents per lb. copper. (Mining and Eng. World, Sept. 6, 1913, 
Vol. 39, No. 10, p. 423, Copper Leaching at Butte, Montana, 
Peter E. Peterson.) 

Bullwhacker Leaching Plant, Butte, Montana. 

The ore is from the same ore body as that of the Butte-Duluth 
Mining Company. The process is : Crush fine. Agitate ore with 
dilute sulphuric acid, using the Hendryx agitator. Separate 
liquor by decantation. Precipitate the copper electrolytically in 
circular vats, without heating, but agitating the electrolyte. 
Remove half the copper by electrolysis, and return the electro¬ 
lyte to storage tanks for further leaching of ore. Current density, 
13 amperes per sq. ft. Extraction, 90 percent. Approximate cost 
per pound of copper, 8 to 10 cents. (Mining and Eng. World, 
Oct. 4, 1913, Vol. 39, No. iq, p. 585, Bullwhacker Leaching 
Plant, Peter E. Peterson.) 

Keith Process at Arlington, N. J. 

This is an electrolytic sulphuric acid process, in which the 
current density at the cathode is maintained proportional to the 
strength of the solution in copper. The tanks cascade 128 in 
series. The size of the tanks and the electrode surface increase 
progressively as the copper content of the electrolyte decreases. 
The current density with 6 percent copper solution is 15 to 20 
amperes per sq. ft. Voltage per cell, 1.6 volts. (Greenawalt, 
p. 318; A. I. E. E., 1902, S. N. Keith.) 


TREATMENT of copper ore. 


235 


Electrolytic Extraction of Copper from Ore at Medzianka, 

Poland, Russia. 

This is an electrolytic sulphuric acid process using anodes 
which are wrapped in permeable envelopes of fabric. The thick¬ 
ness of the fabric is inversely proportional to the current density. 
Anodic oxidation of cations is prevented. Current efficiency, 90 
percent; copper deposited, 0.77 lb. per kw. hour; energy efficiency, 

2 ^ = 36 percent. (Greenawalt, p. 313; U. S. Patent No. 

757,817, April 1, 1904.) 

Plant of Intercolonial Copper Co., N. S., Canada. 

The process is: Roast ore to sulphatize the lime and oxidize 
iron sulphide to ferric oxide. Drop hot ore into 5 percent sul¬ 
phuric acid. Send the copper sulphate solution to storage tank. 
Blow in sulphur dioxide. Send reduced solution, to electrolyzing 
tanks, which are arranged in cascade. Force sulphur dioxide into 
electrolyte through hard rubber tubes, which acts as agitator. 
The anodes were slowly sulphatized, much less rapidly than they 
would have been peroxidized were SO s absent. Current density 
6 amperes per sq., ft. Voltage, 1.5 volts. Electrodes spaced 
iy 2 in. apart. Current efficiency 90 percent. The solution enters 
the electrolyzer with 2.5 percent copper and leaves with 1 percent 
copper. The process is cyclic, continuing indefinitely. Ore con¬ 
tains 2.5 percent of copper, and tailings 0.1 percent. (Greena¬ 
walt, p. 316; Electrochemical Industry, Apr., 1903.) 

Tossizza Process. 

Ferruginous copper sulphate solutions are electrolyzed while 
introducing sulphur dioxide into the electrolyte. The voltage is 
so adjusted that copper is deposited without affecting the iron. 
Voltage 0.2 to 0.6 volt. (Greenawalt, p. 330; U. S. Patent No. 
710,346, Sept. 30, 1902.) 

Siemens-Halske Process. 

The ore is ground fine and roasted at moderate temperature. 
It was the intention of the inventor to so roast that most of the 
iron would be converted to ferric oxide, while most of the copper 



236 


ROBERT RHEA GOODRICH. 


would remain as cuprous sulphide. This operation could not be 
performed. An equally satisfactory roast can be obtained when 
conducted at a moderate temperature (450° C. to 480° C.), most 
of the copper being converted into soluble sulphate. Do not roast 
at a high temperature (dead roast), because much cupric oxide 
would combine with silica to form silicate, and with ferric oxide 
to form ferrite, which are difficultly soluble compounds. The 
active solvent of the process is ferric sulphate, which should 
dissolve all the cuprous sulphide left undecomposed during the 
roasting. In the operation of this process, ferric sulphate becomes 
reduced to ferrous sulphate. In the electrolytic cell the ferrous 
sulphate acts as a depolarizer with regeneration of ferric sul¬ 
phate. A diaphragm is used between anodes and cathodes. See 
equation I (c) ; theoretical voltage = 0.36 volt. In experimental 
tests, impressed voltage used varied from 0.75 to 1.8 volts. This 
process is not in practical use. (Greenawalt, p. 319.) 

Experiments at the Ray Mines, Arizona, 
by W. Y. Westervelt. 

The ore at Ray mines is a great body of disseminated sulphides. 
The process decided upon for the experiment was along the lines 
of the Siemens-Halske process. One ton of 2.5 percent copper 
ore was treated. Extraction, 80 percent; 1.06 lb. Cu was pro¬ 
duced per kw. hour. Cost of copper per pound, 4.5 cents, which 
included only the operating expense. When charged with mining, 
transportation, etc., the cost would be 14.5 cents per pound of 
copper. (Greenawalt, p. 327; Mines and Methods, Oct., 1910, 
W. L. Austin.) 

The Robertson Process. 

The process is: Grind the ore fine. Roast at low temperature 
with the intention of producing a major amount of sulphates. 
Agitate the ore by steam, air or gases under pressure in an 
electrolytic tank with cone-shaped bottom, and use the proper 
electrolyte, usually the sulphate. The copper becomes dissolved 
from the ore at the anode, thereby acting as depolarizer, and is 
precipitated on the cathode. The solution is used repeatedly 
on new lots of ore. Claimed extraction, 90 percent. Current 
efficiency, 70 to 90 percent. Voltage, 1.6 volts. (U. S. Patent 


TREATMENT OF COPPER ORE. 


237 


978,211, Art of Extracting Metals Electrolytically, Dec. 13, 1910, 
James Hart Robertson, N. Y.; U. S'. Patent 988,210, Mar. 28, 

I 9 II v) 

2 . Electrolytic Chloride Processes. 

There are three general classes of processes based on the sol¬ 
vents hydrochloric acid, ferric chloride, cupric chloride. 

The Body Process for the Extraction of Gold, Silver 
and Copper from Ores. 

This process, patented in Belgium in 1883, is of interest 
historically. The ore is ground fine and roasted. It is subjected 
to the action of ferric chloride solution while an electric current 
is simultaneously passing through. The metal becomes dissolved, 
and is precipitated at the cathode. Chlorine liberated at the 
anode reconverts ferrous to ferric salt. Note that, theoretically, 
the efifect of the ferrous chloride depolarizer is to reduce the 
required impressed voltage from 1.35 [called for by equation 
II (a)] to 0.15 volt [that called for by equation II (b)]. (Greena- 
walt, p. 340; U. S. Patent 333,815, Jan., 1886.) 

The Hoepfner Process . 

Cupric chloride, together with sodium or calcium chloride, is 
used as the solvent. The ore is ground fine and treated with the 
cupric chloride solution, which becomes reduced to cuprous 
chloride. The solution is then split in two streams, one passing 
through the anode compartment and the other through the 
cathode compartment. The anode stream acts as a depolarizer 
and becomes regenerated to cupric chloride, while the copper is 
precipitated from the cathode stream. On issuing from the 
electrolyzer the two streams are reunited and constitute the regen¬ 
erated solvent, containing one-half the copper contents of the 
original stream. A diaphragm separates the anode and cathode 
compartments. Note that, theoretically, the effect of the cuprous 
chloride depolarizer is to reduce the required impressed voltage 
from 1.53 [called for by equation III (a)] to 0.84 volt [that 
called for by equation III (b)]. It is claimed for this process 
that practically 0.6 to 0.8 volt is required. This process was used 
at Schwartzenberger Hiitte, in Saxony, from August, 1891, to 


238 ROBERT RHEA GOODRICH. 

March, 1892. The results were unsatisfactory. The main difficulty 
appears to be metallurgical, the cupric chloride being a somewhat 
indifferent solvent. The question arises, “Would this process 
not work better on roasted ore?” (Greenawalt. p. 342). 

Leaching Experiment on Keystone Ore at Miami, Arizona. 

The ore contains 3 to 5 percent of copper, occurring on the 
surface in veinlets of chrysocolla (CuSiO s + 2H 2 0) in white 
porphyry. It was desired to employ a process in which the 
solvent was regenerated and which produced marketable copper. 
The process used for experiment was a modification of the 
Hoepfner process. The ore crushed to' in. (2 cm.) was first 
treated in a kiln, wherein it was subjected to the action of reducing 
gases from a small gas producer. The reduction was fairly com¬ 
plete, almost all the copper appearing in a metallic state in the 
calcined product. The lixiviant contained 5 percent of copper 
as cupric chloride, and 20 percent of sodium chloride. Calcined 
ore, 12 mesh, treated with hot leaching solution for 2 hours, gave 
83 percent extraction; 60 mesh, 99 percent extraction. Electric 
light carbons were used for anodes. The potential employed per 
cell was 1 volt, with current density of 10 to' 12 amperes per 
sq. ft. (1.1 to 1.32 per sq. dm.). One ton of copper was pro¬ 
duced. The final washings from the leaching barrels were run 
over scrap iron, and the cement copper was used to' reduce the 
last of the cupric chloride to cuprous chloride in the lixivium 
before going to the electrolytic cell. (Mines and Methods, Nov., 
1911, P- 355 , Leaching Applied to Copper Ore, W. L. Austin). 

The Greenawalt Process. 

The copper in the ore is dissolved by dilute acid chloride solu¬ 
tion and precipitated by electrolysis. The acid is regenerated and 
augmented at the expense of sulphur dioxide and water. In 
practice, there is required to be supplied 2 ounces (58 grams) of 
salt per pound of copper extracted. Since the copper in the 
electrolyzer is in the cuprous state, twice the amount is deposited 
for a given amperage. The process may be applied to the fol¬ 
lowing ores: 1. Siliceous oxidized copper ores. 2. Siliceous 
sulphide copper ores. 3. Copper concentrates. 4. Siliceous gold 
and silver ores containing copper. 


treatment oe copper ore. 


239 


Greenawalt claims that for treatment of No. 4 ores there is no 
other satisfactory process. If there is much iron in the ore, it is 
desirable to roast. Sulphide ores necessitate roasting. In treat¬ 
ing these various ores experimentally an extraction of 90 to 98 
percent of total copper, gold and silver contents has been obtained. 
Theoretical impressed voltage was 1.53 volts. (See equation 
III (a).) Copper is deposited 1 lb. per lew. hour. The esti¬ 
mated cost of producing copper is 2.7 cents per pound, adminis¬ 
tration expense not included. (Greenawalt, p. 349; E. and M. 
Journal, Nov. 26, 1910; U. S. Patent, 673,776, Oct. 25, 1910.) 
The Recovery of the Gold and Silver Contents of Copper Ores. 

Most copper ores treated by wet methods contain valuable 
amounts of gold and silver, the recovery o-f which is important 
in many cases. While the assay returns may be practically nil, 
the aggregate recovery in a year may be considerable. Utah 
Copper Company during the year 1910 recovered on an average 
23 cents value in gold and silver per ton of ore treated. 

The gold and silver may be recovered: (a) Before the copper 
is precipitated; (b) simultaneously with the copper; (c) by a 
subsequent operation. 

(a) With the Hunt and Douglas process, when the leaching 
is done with ferrous chloride and salt solution, silver is converted 
into AgCl by CuCl 2 . After reducing all copper to the cuprous 
condition the silver may be precipitated on copper. 

(a) and (b) In Greenawalt’s U. S. Patent 973,776, when 
there is gold in the ore he increases the free chlorine in the 
lixiviant, and thus both gold and silver are dissolved. Then 
leaching with fairly concentrated solution of base metal chlorides, 
dissolving both the gold and silver with the copper, the gold 
and silver may be precipitated prior to, or with the copper, by 
varying the current density. 

(a) and (b) In the Bradley process the gold and silver are 
brought into solution, and may be precipitated with the copper, 
or separately. 

(b) The Longmaid-Henderson process, while formerly pre¬ 
cipitating the silver prior to copper, now precipitates them 
jointly, since the electrolytic refineries pay for 95 percent of 
silver and all of the gold and copper. 


240 


ROBERT RHEA GOODRICH. 


(c) In the Hunt and Douglas process, when leaching roasted 
ore with dilute sulphuric acid, the gold and silver are left in the 
residue. Silver is extracted by brine, and gold by chlorination, 
amalgamation or cyanidation. (Mines and Methods, Mar., 19 12 , 
p. 433, Leaching Applied to Copper Ore, W. L. Austin.) 

Summarizing and Looking Toward the Future. 

A number of important copper mining companies are experi¬ 
menting with the development of hydro-metallurgical methods. 
Mr. Laist, at Anaconda, expects to produce copper for 6.5 cents 
per pound from concentrator tailings. There are two concerns in 
Butte, Montana, The Butte-Duluth Company and The Bull- 
whacker Company, which are producing finished copper on a 
commercial basis. Throughout the Southwest experimenting is 
being conducted. The Arizona Copper Company, at Clifton, 
Arizona, has been using leaching methods for treating concen¬ 
trator tailings of their oxide mill for many years. The recent 
experiments of the Shannon Copper Company, at Clifton, Ari¬ 
zona, with a view to reducing acid consumption on an ore with 
a basic gangue, are very encouraging. The Braden Copper Com¬ 
pany, Chile, is experimenting, leaching one-fifth of their output 
of concentrates. The Chile Exploration Company is extracting 
copper from an ore which is very amenable to< leaching. Spain, 
Germany, England and Russia have commercial plants or are 
experimenting in this field. 

It is said there is hardly a known chemical reaction which has 
not been applied. It would appear that no method of universal 
application is likely to be developed, that present methods will 
be perfected, and that the one most suitable will be selected for 
the particular case. 

Passing the different methods of extraction in review, we may 
generalize: 

I. Dissolving the copper. 

II. Precipitating the copper. 

I. (a) In completely oxidized ores, by crushing and leach¬ 
ing with dilute acid direct. 

(b) In a mixture of oxidized and sulphide ores, by 


TREATMENT OF COPPER ORE- 


241 


sulphatizing or chloridizing, roasting, and leaching with dilute 
acid. 

(c) In straight sulphide ores, by sulphatizing or chlorid¬ 
izing, roasting, and leaching with acid. 

(d) In ores, by a modification of (a), (b) and (c),« 
taking advantage of the solvent action of solutions containing 
ferric salts. 

II. (a) Precipitating by chemical reagents. 

1. Iron—cast iron, sponge iron, scrap iron. 

2. Lime. 

(b) By electrolytic deposition, with insoluble anodes of 
graphitized carbon for chloride solutions; and peroxidized lead 
anodes, or magnetic oxide anodes for sulphate solutions. (Fe 3 0 4 
anode, made by fusing magnetite, and casting.) 

The precipitation of the copper will be: 

1. From pure or purified electrolyte (one free from iron 
salts). 

2. From ordinary impure electrolyte, using sulphur dioxide 
as depolarizer without a diaphragm, and securing (a) regen¬ 
eration of acid solvent and (b) high current efficiency and high 
energy efficiency. 

3. From ordinary impure electrolyte, using ferrous salt with 
a diaphragm as depolarizer, regenerating the ferric salt as a 
solvent, and securing high current efficiency and high energy 
efficiency. 

There are two important economic considerations which alone 
should goad the metallurgist to future discoveries in hydro- 
metallurgical methods: 

1. Enhancement of the life and value of a property 5 ° P er “ 
cent when the hydro-metallurgical process extracts 90 percent, as 
against 60 percent by the combined concentrating-smelting proc¬ 
ess, assuming equal cost per pound of metal produced in the 
two cases. 

2. Possible treatment at a profit of certain complex mineral- 


242 


ROBERT RHEA GOODRICH. 


aggregates not otherwise amenable to extraction, recovering 
the several metals. 

Articles not previously accredited: W. L. Austin’s numerous 
consecutive articles in Mines and Methods, Sept., 1910, to Oct., 
1912; E. and M. Journal, Oct. 4, 1913, Vol, 96, No. 14, p. 651, 
Leaching of Copper Ores; E. and M. Journal, Nov. 22, 1913, 
Vol. 96, No. 21, p. 962, Copper Leaching; Mining World, May 
17, 1913, Vol. 38, No. 21, p. 947, Baxeres de Alzugaray, Exten¬ 
sion of Hydro-metallurgical Industries; Mining and Scientific 
Press, July 19, 1913, Vol. 107, No. 3, p. 127; Mining and 
Scientific Press, Aug. 16, 1913, Vol. 107, No. 7, p. 252, Sul¬ 
phuric Acid Leaching; Metallurgical and Chemical Engineering, 
1913, p- 600, Leaching Copper Ores and Tailings. 

Credit is also due to Dr. Edward F. Kern, of the Metallurgical 
Department of Columbia University, for kindly looking over the 
manuscript and offering suggestions. 


PHOTOGRAPHS. 


Gi—Photo-micrograph. 
Ga—Photo-micrograph. 
G 3 XN—Photo-micrograph. 
No. i—Testing Table. 

No. 2—Electrolytic Cells. 


Pages 37 to 45. 

No. 3—Frame. 

No. 4—Cell Assembled. 
No. 5—Cell Taken Apart. 
No. 6—Electrodes. 


TABLES. 

Pages 46 to 78. 

No. 1—Screen Analysis of Crushed Ore, Basis of Curve Sheet No. 1. 

No. 2—Roasting in the Gas Muffle Furnace: 

2-A, Roasting at 500° C.; 

2-B, Roasting at 600 0 C.; 

2-C, Roasting at 725 0 C.; 

2-D, Roasting at 850° C.; 

2- E, Basis of Curve Sheet No. 2. 

No. 3—Roasting in the Large Gas Furnace: 

3- A, B-Series, through 20 on 40-mesh ; 

3-B, B-Series, through 40 on 80-mesh ; 

3-C, B-Series, through 80-mesh ; 

3-D, B-Series, whole through 20-mesh ; 

3 -E, A-Series; 

3-F, Comparison. 

No. 4—Leaching in Bottles, Basis of Curve Sheet No. 4. 


1 Basis of Curve Sheets 
f Nos. 3-A, 3-B, 3-C, 
) 3 -D. 


No. 5—Test No. 3: 

5-A, Copper Refining; 
5-B, Copper Refining; 
5-C, Anodes; 

5-D, Depolarization ; 
5-E, Depolarization. 


* Basis of Curve Sheet No. 5. 


y 


No. 6—Tests Nos. 1, 2, 4: 

6-A, Preliminary Tests; 

6-B, Tests Nos. 1, 2, 4 ; 

6-C, Tests Nos. 2, 4 (Finishing Step), Basis of Curve Sheet No. 6. 


Examples of Step. Arrangement of Plant: 
Arrangement A, 

Arrangement B, 

Arrangement C, 

Arrangement D. 




CURVE SHEETS. 

Pages 79 to 87. 

No. 1—Screen Analysis of Crushed Ore. 

No. 2—Roasting in the Gas Muffle Furnace. 

No. 3—Roasting in the Large Gas Furnace: 
3-A, Through 20 on 40-mesh; 

3-B, Through 40 on 80-mesh; 

3-C, Through 80-mesh; 

3-D, Through 20-mesh. 

No. 4—Leaching in Bottles. 

No. 5—Test No. 3. 

No. 6—Finishing Step: 

Test No. 2; 

Test No. 4. 

DIAGRAMS. 

Pages 88 to 91. 

Examples of Step Arrangement of Plant: 
Arrangement A, 

Arrangement B, 

Arrangement C, 

Arrangement D. 


EXPERIMENTS ON A PORPHYRY COPPER ORE 
FROM BISBEE, ARIZONA. 

Ibis report is given under the following heads: Petrographic 
Description (Microscopic Study) ; Sampling and Preparing Ore 
for Treatment; Chemical Analysis; Treatment of Ore. 
Treatment of Ore: 

1. Oxidizing Roast; 

2. Leaching with Dilute Sulphuric Acid; 

3. The Electrolytic Precipitation, 

(a) The Electrolytic Plant, 

(b) Tests Nos. 1, 2, 3, 4, 

(c) Step System of Electrolytic Precipitation. 

PETROGRAPHIC DESCRIPTION. 

Microscopic Study for Classification 

Texture : Variable. Fine to medium mixed aggregate. 

Original Structure : Obscure. 

Secondary Structure : Fractured ; healed. 


MINERALOGY (Minerals are grouped for interpretation purposes and are arranged in each 

group in approximate order of abundance) 


PRIMARY 

Essential Minerals 

PRIMARY 
Accessory Minerals 

SECONDARY 
Alteration Products 

Obscure 

Probably Quartz, 
and Feldspar 
now wholly altered 

Zircon 

Quartz 

Kaolin 

Sericite 

Introduced Substances or 
Mineralization 

Yellow Sulphide (Pyrite) 

Black Sulphide (Chalcocite) 

Quartz 

Tertiary Changes and Enrichment 
Effects on Ores 

Green Malachite 

Limonite (a little from the Yellow 
Sulphide) 


The original character of the rock has been much obscured 
by modification. The variable texture indicates that the original 


5 













6 


ROBERT RHEA GOODRICH. 


rock was much brecciated. Silicification seems to have been a 
prominent feature. Some of the quartz appear to be remnants 
of primary grains. It must have been primarily an acid intru¬ 
sive, which, after having been fractured and brecciated, has 
become still more acid by silicification. There appears to be 
recorded several sets of movements. The special features of 
most importance seem to be as follows: 

The Ground Mass. 

This shows variable texture, fragmental (brecciated). The 
whole is so much modified that the fragmental character is not 
plain. Most of the grains seem to be quartz. The only evi¬ 
dence of feldspar are those areas judged to be slightly kaolin- 
ized and areas of sericite. The fragmental character of the 
rock is somewhat further emphasized by the manner in which 
the metallics are distributed. These are distributed in such a 
way as to suggest an original fragmental (breccia), into the 
interstices of which were introduced the yellow and black sul¬ 
phides. Certain streaked areas are much clearer than the rest 
of the ground mass and generally free from introduced metallics. 
These are filled with what is plainly introduced quartz, and they 
seem to represent old fractures. It is difficult to say just what 
relation these bear to the mineralization periods; they may pos¬ 
sibly represent the closing stages of silicification. 

The Introduced Metallics. 

These consist of yellow pyrite and black chalcoeite. The 
yellow sulphide seems to be a little too yellow for typical 
pyrite and not quite yellow enough for chalcopyrite. The yel¬ 
low sulphide bears evidence of having been fractured and some¬ 
what crushed. The fractures are healed with the black sulphide. 
The fractures extending across the grains of yellow sulphide 
end abruptly at their margins, and do not continue into the 
adjacent grains of ground mass. Under the microscope, there 
may be seen areas of yellow sulphide, veined and rimmed with 
black sulphide; the whole is rimmed with a narrow band of 
sericite and, in some cases, of quartz. It is evident that the 
history of this rock is very complicated. 


TREATMENT OF COPPER ORES. 


7 


It would seem, therefore, that two periods of mineralization 
are here represented: first, the decomposition of the yellow sul¬ 
phide, followed by sufficient movement to cause fracturing and 
slight crushing, either during the closing stages of its deposi¬ 
tion or immediately thereafter; second, the introduction of the 
black sulphide, which was deposited in the previously formed 
and fractured yqjlow sulphide and as an envelop surrounding 
the smaller grains of the yellow sulphide. Silicification very 
likely accompanied all of these changes. According to this in¬ 
terpretation, the rock is a silicified and mineralized brecciated, 
acid intrusive (quartz porphyry), in which two periods of min¬ 
eralization are represented. There is no absolute proof in these 
slides as to the source of the secondary mineralization, i. e., 
the chalcocite.* 

Photo-micrographs, G 1} G 2 , G 2 XN. (Magnification go diameters.) 

G x —The large patches, light in shade, comprising the greater 
part of the photo-micrographs, are yellow pyrite. There are two 
large patches, and several smaller ones, light in shade, but differ¬ 
ing in appearance from those representing pyrite, in that they 
have a surface of homogeneous shade. These are quartz. The 
darker portions, occurring as spots and veinlets cutting the yellow 
pyrite, are of chalcocite. 

G 2 —In this photo-micrograph, the black masses and veinlets 
of chalcocite are very strongly marked. There is also some 
white non-metallic material. 

G 3 XN—This photo-micrograph was taken with crossed nicols 
of a slide containing more non-metallic material. The yellow 
pyrite here shows as large patches, quite dark, intersected by 
black veinlets of chalcocite. There is much white non-metallic 
material which, being doubly refracting, shows white as in the 
other photo-micrograph (taken with upper nicols out), while the 
pyrite, not being doubly refracting, shows darker. 

SAMPLING AND PREPARING ORE FOR TREATMENT. 

The ore was received in lump form. Five hundred pounds 
were crushed to pass a 4-mesh screen, using the gyratory crusher 

* This abstract was taken from the petrographic description by R. J. Colony, made 
under the direction of Dr. Charles P. Berkey, Columbia University. 


8 


ROBERT RHEA GOODRICH. 


followed by the cone and ring sample grinder. A sample for 
analysis was cut out and ground through ioo-mesh. One-quar¬ 
ter of the lot was then cut out for treatment. This was passed 
repeatedly through the sample grinder (which was set up so 
as to grind finer), until all passed a 20-mesh screen. This 
one-quarter was then thoroughly mixed by repeated coning, and 
then split by split shovel. One resulting one-eighth was re¬ 
tained for treatment and will be called “Whole through 20- 
mesh.” The other resulting one-eighth was sized, producing 
the portions: Through 20 on 40-mesh, 45.4 percent of whole; 
through 40 on 80-mesh, 22.9 percent of whole; through 80- 
mesh, 31.7 percent of whole. (The screen analysis of the 
crushed ore is given in Table No. 1 and in Curve Sheet No. 1.) 


CHEMICAL ANALYSIS. 

Percent 


Si 0 2 (insoluble) . 65.10 

Fe . 10.90 

CaO . 0.90 

AUO3 . 0.28 

Total Cu . 6.04 

S . 12.70 


1.70 Sol. in dil. 
HC 1 . 

4.34 Bal. of Cu. 


TREATMENT OF ORE. 

It has been seen that the ore is mainly a sulphide. Copper 
sulphide ores, which contain an excess of silica, have been here¬ 
tofore usually treated by mechanical concentration followed by 
smelting. But this ore, like much of the disseminated ore bodies 
of the Southwest, contains a portion of its valuable copper 
contents in an oxidized condition—malachite. Mechanical con¬ 
centration on pure sulphides makes usually a saving of 66 per¬ 
cent ; and, when there is an oxidized component, the saving is 
likely to be still lower. For this reason, it was decided to make 
a test on this ore by a leaching method. The method selected 
comprises: 

1. Oxidizing Roast; 

2. Leaching with Dilute Sulphuric Acid; 

3. Electrolytic Precipitation of the Dissolved Copper. 








TREATMENT OF COPPER ORES. 


9 


Oxidizing Roast. 

A series of four roasts was made in a gas muffle furnace. 
(17^2 in. by nj/2 in. by 4 in.—inside measurement.) In order 
that the relative roasting qualities of coarse and fine material, 
as well as the most suitable temperature, might be determined, 
300 grams of through 20 on 40-mesh material, and 300 grams 
of through 80-mesh material were roasted, in two seven-inch 
roasting dishes. The temperature of roasting was measured 
by a thermo-electric pyrometer. Samples to determine the 
progress of roasting were taken at half-hour intervals. These 
samples were ground through 100-mesh and then tested as fol¬ 
lows : One gram of sample was boiled in a covered caserole for 
twenty minutes, with 150 c. c. of water. The soluble copper 
thus obtained is reported in percent copper soluble in water. 
After filtering and washing, the same portion of sample was 
boiled in a covered caserole for twenty minutes, with 150 c. c. 
of dilute hydrochloric acid (100 c. c. of concentrated hydrochloric 
acid diluted to 1,000 c. c.). The soluble copper thus obtained 
is reported as percent copper soluble in dilute hydrochloric acid. 
The sum of these two is reported as the percent of total soluble 
copper. The difference between this and the total percent of 
copper found in the sample, is reported as percent insoluble 
copper. By dividing percent of total soluble .copper by total 
percent of copper in sample, percent extraction was determined. 
(See Tables Nos. 2-A, 2-B, 2-C, 2-D, 2-E.) 

On referring to Curve Sheet No. 2, Roasting in the Gas Muffle 
Furnace, where material through 20 on 40-mesh, and material 
through 80-mesh were roasted side by side in separate roasting 
dishes, it will be seen, by leaching with dilute acid, that the 
material through 20 on 40-mesh always gave a higher extrac¬ 
tion than the material through 80-mesh. As regards temper¬ 
ature, the roast conducted at 850° C. gave the poorest extrac¬ 
tion, the temperature being too high. (The series of roasts 
indicates that at temperatures above 725 0 C. the resulting cop¬ 
per oxide forms insoluble compounds.) The roast conducted 
at 500° C. gave better extraction, but as shown by subsequent 
roasts, the temperature was too low to secure the best extraction 
on leaching. Roasts conducted at 600 3 C. and 7 2 5 ° C. are 


10 


ROBERT RHEA GOODRICH. 


equally good on material through 20 on 40-mesh; but on mate¬ 
rial through 80-mesh, the 6oo° temperature gave material which 
yielded the best extraction. Consequently the temperature of 
6oo° C. was selected as the best for subsequent roasting in the 
large gas roasting furnace. 

There was no muffle in the large gas roasting furnace and its 
operation was similar to that of a gas-fired reverberatory fur¬ 
nace. The hearth, which was removable, consisted of a sheet- 
iron pan, lined with fire brick. (22^ in. by 9 in.—inside meas¬ 
urement.) Two series of roasts were made; viz., A-Series and 
B-Series. Eight pounds of material were roasted per charge. 
The temperature of 6oo° C. was held constant. The materials 
roasted were: through 20 on 40-mesh; through 40 on 80-mesh; 
through 80-mesh; whole through 20-mesh. In all about 100 
pounds of ore were roasted. Samples were taken during roast¬ 
ing at quarter-hour intervals. Of the several roasts made of 
one size material, in the A-Series and in the B-Series as through 
20 on 40-mesh, average samples were made. These samples 
were ground through 100 mesh, and tested for soluble copper, 
in the same manner as the samples resulting from the roasting 
in the gas muffle furnace. The comparison of total soluble cop¬ 
per determined by using dilute hydrochloric acid with that de¬ 
termined by using dilute sulphuric acid, may be seen in Table 
No. 3-F. 

Referring to Tables 3-A, 3-B, 3-C, 3-D, and Curve Sheets 
Nos. 3-A, 3-B, 3-C, 3-D, Roasting in the Large Gas Furnace, 
it will be seen that an extraction of 96 percent was secured in 
all roasts, except on the through 20 on 40-mesh material, which 
was but 93 percent. The roasts were all carried on for a longer 
time than was necessary. An inspection of the curves will 
enable one to determine the most desirable time of drawing the 
roasts. (See Tables 2-A, 2-B, 2-C, 2-D. 2-F, 3-A, 3-B, 3-C, 
3~D, 3-F, 3-F; Curve Sheets Nos. 2, 3-A, 3-B, 3-C, 3-D.) 

Leaching with Dilute Sulphuric Acid. 

In order to forecast, as well as could be done in a small way, 
what might be expected in practice, regarding extraction and 
acid consumption, the following leaching experiments were 
made: 20 grams of roasted ore, together with 200 c. c. of 10 


TREATMENT OF COPPER ORES. 


11 

percent H 2 S 0 4 ,* were introduced into glass stoppered bottles, 
which were continually shaken. Two tests were made, one at 
a temperature of 21 0 C.; the other at a temperature of ioo° C. 
The material here leached was in the same condition, as regards 
degree of comminution, as when it left the sizing seive, except 
as the roasting may have affected it. It is to be noted that it 
was not so with the samples of roasted materials earlier tested 
for solubility of the copper contents, and which were all ground 
through ioo-mesh before boiling in the caserole with water and 
dilute hydrochloric acid. 

In the test made at 21" C., the extraction ranged from 72 
to 90 percent. The extraction was not improved by continuing 
the operation longer than 6 hours. The acid consumption was 
moderate, being 2.0 pounds of Oil of Vitriol (66° Baume) per 
pound of copper extracted, as compared with 1.65 pounds theo¬ 
retically required to dissolve one pound of copper existing as 
oxide. The test which was made at ioo° C. gave an extraction 
of 90 percent with the through 20 on 40-mesh material, while 
with all other materials the extraction was 100 percent. The 
acid consumption was somewhat higher as Curve Sheet No. 4 
shows. The time required for efficient leaching was between 
three and six hours. Nothing was gained by increasing the 
time of leaching beyond six hours. (See Table No. 4; Curve 
Sheet No. 4.) 

The Electrolytic Precipitation. 

The Electrolytic Plant. 

The electrolytic plant, designed for experimenting in the de¬ 
position of copper from solution with insoluble anodes, com¬ 
prised a motor generator set, circulating pumps, electrolytic cells, 
and a testing table. (See Photos Nos. 1, 2, 3.) 

The motor generator set used was a Robbins and Meyers 
34 HP, no volt, 60 cycle, 1,750 rpm., single phase motor, 
belted to their 0.125 kw., 10 volt, D. C., 1,750 rpm., compound 
wound generator. 

A small acid-proof pump could not be found in the market, 
so a single-acting acid-proof pump was designed. Four pumps 

* In this paper percent of a constituent of a solution signifies grams of the con¬ 
stituent per 100 c.c. of the solution: “percent IL.S0 4 ” signifies grams of free ILSO4 
per 100 c.c.; “percent copper,” grams of copper per ioo c.c. 


12 


ROBERT RHEA GOODRICH. 


were made. A 2-oz. syringe, No. 135, made by the American 
Hard Rubber Company, was used as a pump barrel. A block 
of dry maple was suitably bored. There were two ball valves; 
the seats were rings of ^4-in. hard rubber; the balls, glass agates. 
The balls were ground into the seats with emery. The seats 
entered the cavities bored for them snugly. P. and B. acid-proof 
paint secured the seats in place so that there was no leakage. 
The syringe, the suction pipe and the discharge pipe, entered 
holes in the maple block prepared for them. Water-tight con¬ 
nections were made by suitable soft-rubber packing rings. The 
Robbins and Meyers, ps HP, no volt, 60 cycle, 1 , 75 ° r P m -, 
single phase motor, through suitable pulleys and gears operated 
the four pumps, which made 35 strokes per minute. (See PhotO' 

No - 3 -) 

The electrolytic cells used in the experiments of this paper 
were 4 cells with 1 cathode each, 2 cells with 2 cathodes each, 
and 2 cells with 4 cathodes each. Each cell had one more anode 
than cathodes. (See Photo No. 2.) Photographs Nos. 4 and 5 
show one cell with 4 cathodes, assembled and taken apart. The 
electrodes were spaced Tt in* from center to center. Anode No. 
6, seen in Photograph No. 6 (lower left hand corner), was used 
in Tests Nos. 1 and 2. It was made from a sheet of lead (8 lb. 
to the sq. ft.), 2p> in. wide and of the proper length. The upper 
edge was wrapped around a piece of No. 6 bare copper wire 
by which it was suspended in the cell. Cathodes may likewise 
be seen in Photo No. 6 (middle row). The cathode was made 
of a sheet of thin copper attached to’ a copper clamp by which 
it was suspended in the cell. There were four insulating blocks 
of hard rubber, 1 in. by 1 in. by ^ in., beneath the cell. There 
were four pieces of hard rubber, 1 in. by 1 in. by T 5 ¥ in., with a 
i 3 6 in. upward projection for insulating and holding the bus¬ 
bars in place. In Photograph No. 5, just below the cell, the 
bus-bars may be seen. The bus-bar was comprised of a 1 in. 
by 1 in. brass angle and-a 1 in. by pj in. hard rubber strip. 
The rubber strip was bolted to' the vertical leg of the brass angle. 
There were two pieces of hard rubber spacing-pieces used for 
holding the bus-bars rigidly in place. The method of connect¬ 
ing the electrodes with'the bus-bars was similar to practice. 
The brass and the hard rubber of the bus-bars were so cut that 


TREATMENT OF COPPER ORES. 


13 


when the horizontal arms of the electrodes rested on them, on 
one side the arms of the anodes rested on the brass plus con¬ 
ductor while the arms of the cathodes were insulated by hard 
rubber, and on the other side the arms of the anodes were insu¬ 
lated by hard rubber while the arms of the cathodes rested on the 
brass minus conductor. The binding posts with No. 6 bare copper 
wire served for connecting between cells. (See Photos Nos. 
4 > 5 -) 

Quite elaborate piping systems about the cells may be seen 
in Photo No. 2. The elevated temperature tank and the cells 
of a step were connected by i-in. lead pipe, the connections to 
and from the step circulating pump were made by y 2 - in. lead 
pipe. For the progressive circulation of the electrolyte through 
the plant, y 2 - in. lead pipe was used. The sulphur dioxide gas 
was conveyed to the cells by its system of lead piping. The gas 
was supplied from a cylinder of liquid gas. The gas cylinder 
was connected by small rubber tubing to the end of a 2-ft. length 
of i-in. lead pipe. From the rear side of this pipe, four V 2 - in. 
lead pipes, leading to different parts of the plant, passed beneath 
the cells. (See Photo No. 2, in foreground.) There was placed 
horizontally above the cells of each step a pAin. lead pipe. This 
pipe was fitted with y~in. tee connections which were connected 
to the hollow anodes, used in Tests Nos. 3 and 4, by rubber 
tubing. Bottles were introduced between the ends of the four 
distributing pipes and these horizontal pipes, just mentioned, in 
order to indicate the amount of gas flowing. The four indicator 
bottles for judging the rate of flow of the gas contained a little 
water through which the gas bubbled. A screw clamp on the 
rubber tubing connection, between each distributing pipe and its 
indicator bottle, adjusted the distribution of gas. (See Photo 
No. 2.) 

A poplar kitchen table, with top 30 in. by 48 in., was selected 
for the testing table. A Weston miniature, precision, direct- 
current volt-ammeter was used (Model 280 Triple Range Port¬ 
able Volt-ammeter, Weston Electrical Instrument Co., Newark, 
N. J., Bulletin No. 8, 1912.) The scales were: 150, 15, 3 volts; 
30, 15, 3 amperes. The instrument was permanently placed and 
connected on the table top at the operator’s left hand. The 
table was wired and connected with the rest of the plant, so 


H 


ROBERT RHEA GOODRICH. 


that all the readings desired might be taken on the one instru¬ 
ment when the switches were properly manipulated. The fol¬ 
lowing readings could be taken : 

1. Total amperes, load on generator. 

2. Amperes, No. i load circuit. (Plant could be operated 

as one circuit or as two parallel circuits.) 

3. Amperes, No. 2 load circuit. (Plant could be operated 

as one circuit or as two parallel circuits.) 

4. Generator volts. 

5. Individual volts across cells. 

There were fourteen voltage wires, each terminating at one 
end (the cell end) in a tee connector. In the foreground of 
Photograph No. 2 may be seen the cell end of one voltage wire 
(No. 12). In connecting up the cells (arrangements might be 
varied in different tests), there was always to be found close 
to the cell the cell end of a voltage wire. In making series con¬ 
nections between consecutive cells, No. 6 copper connecting 
wires from the respective binding posts of the bus-bars of the 
two cells entered a tee connector, one leg of which was per¬ 
manently attached to the end (cell end) of a voltage wire. All 
the fourteen voltage wires passed to the testing table where 
they connected to two 13-point circular switches. The switch 
to the operator’s left had connected to it wires Nos. 1 to 13, 
while the switch to his right had connected to it wires Nos. 2 
to 14. On properly setting the two 13-point switches and throw¬ 
ing the 3-pole double throw switch adjoining, the individual volts 
across cells, or the sum voltage of any number of cells, could be re¬ 
corded by the instrument on suitable scale. Not all the cell ends 
of the voltage wires were connected in any one test. When the 
plant was wired up for some one test, the voltage wires in use 
were noted and the corresponding switch points on the table 
top were tagged for the operator's use. All switch points con¬ 
nected with unused voltage wires were dead. 

The generator field rheostat is seen in the foreground of 
Photograph No. 1. To the operator’s left, on the vertical panel, 
were four battery rheostats, connected in parallel, serving as a 
variable load in series with the cells in load circuit No. 2. These 
battery rheostats could be short circuited, when the generator 


TREATMENT OE COPPER ORES. 


15 


field rheostat would give the desired voltage regulation. When 
a lower voltage was desired than the field rheostat could give, 
then the variable resistance in series was used. (See Photos 
Nos. 1, 2, 3, 4, 5, 6.) 

Test No. 1. 

Arrangement A, in “Example of Step Arrangement of Plant/’ 
was employed, operating No. 1 Step and No. 2 Step and omitting 
the Finishing Step. No. 1 Step had four cells, connected in 
series, each with one cathode. No. 2 Step had two cells, con¬ 
nected in parallel, each with two cathodes: this combination will 
be spoken of as one cell of four cathodes. Note that the plant 
was operated with one-third as many cells as are indicated in 
the diagram (Example of Step Arrangement of Plant -A). 
Evaporation was compensated for by adding sufficient water to 
the temperature tank of No. 1 Step and to the temperature tank 
of No. 2 Step. The relative amounts of water for each step 
were proportional to the step surface exposed to evaporation. 
Consequently the overflow to the sump of No. 2 Step was equal 
in volume to that of the copper solution fed to the temperature 
tank of No. 1 Step. '(See Table No. 6-B.) 

Current density of 20 amperes per square foot was selected 
for No. 1 Step. The area (sum of the two sides) of the cathode 
is 0.0824 sq. ft., therefore a current of 1.648 was supplied. The 
volume of copper feed solution required for 12 hours, in order 
to supply the cells of No. 1 Step with an amount of copper equal 
to that deposited in that step by the current selected plus the 
copper carried on by the progressive circulation so that the elec- 
trolyte would remain constant in composition, was determined 
as follows: 

1.648 X 4 x 12 X 1.1855 X 0.90 r ’ 8 7f c -. c - Copper feed 

-- solution required tor 

0.75X 0.06 12 hours. 


Where, 

Current amperes . 1.648 

Number of cells in series, No. 1 Step . 4 

Duration—hours . 12 






i6 


ROBERT RHEA GOODRICH. 


Theoretical copper deposited by i ampere hour— 


grams . 1.1855 

Current efficiency percent, assumed . 90. 

Copper deposited in No. 1 Step, expressed as per¬ 
cent of that in feed . 75. 

Copper in feed solution percent . 6. 


1,872 c. c. of No. 1 Step electrolyte, containing 1.5 percent 
copper, overflowed from the last cell of No. 1 Step to the tem¬ 
perature tank of No. 2 Step. This supplied No. 2 Step with 
an amount of copper equal to that deposited by the current in 
that step, plus the copper carried on to the sump by the pro¬ 
gressive circulation, as the following calculation shows: 


1.648 X 1 X 12 X 1.1855 X 0.90 
0.75 X 0.015 


1,872 c. c., overflow of No. 
1 Step electrolyte, re¬ 
quired for No. 2 Step. 


Where, 

Number of cells in series, No. 2 Step . 1 

Copper in overflow of No. 1 Step, electrolyte percent. . 1.5 


1 he strength of the electrolyte in H 2 S 0 4 , which is developed 
by the acid liberated by the copper on deposition, was deter¬ 
mined as follows: 


No. 1 Step. 

0.06 X 0.75 X i-54 — 0.069 (or 6.93 percent). 

Where, 

Copper in feed solution percent .. . 6. 

Copper deposited in No. 1 Step, expressed as percent 

of that in feed . 73. 

H 2 S 0 4 freed per unit of copper deposited . 1.54 


No. 2 Step. 

0.06X 0.9375 X 1-54 — 0.0866 (or 8.66 percent). 
Where, 

Copper deposited in Nos. 1 and 2 Steps, expressed 
as percent of that in feed. 


9375 















J 


treatment of copper ores. 17 

In starting- up the plant, each step was supplied with the 
necessary volume of electrolyte to put it in operation. The com¬ 
position of this initial charge of electrolyte in copper and acid 
was that which the above calculations indicate and which may 
be seen in Table No. 6-B. The results of this test show that, 
when employing the Step Arrangement of Plant, and when 
adding water to each step to compensate for evaporation, the 
initial values of copper and acid contents of the electrolyte re¬ 
main constant during the electrolysis. The results of this test 
also show the performance when copper is deposited, using in¬ 
soluble lead anodes. (See Photo No. 2; Tables Nos. 6-A, 
6-B; Table and Diagram Example of Step Arrangement of 
Plant-A.) 

Test No. 2. 

Arrangement A, in “Example of Step Arrangement of Plant,” 
was employed, using No. 1 Step, No. 2 Step, and the Finishing 
Step. The arrangement of the cells in Nos. 1 and 2 Steps was 
the same as in Test No. 1. The Finishing Step had two cells, 
connected in parallel, each with 4 cathodes: this combination 
will be spoken of as one cell of 8 cathodes. The operation of 
the Finishing Step of this test will be discussed in Test No. 4. 

In order to simplify the operation, a weaker copper feed solu¬ 
tion was made for No. 1 Step by adding 1,170 c. c. water to 
1,872 c. c. of 6.0 percent copper solution, producing 3,042 c. c. 
of feed solution containing 3.7 percent copper, which was used 
for the 12-hour run. This supplied No. 1 Step with the same 
amount of copper as in the previous test and with the water 
necessary to compensate for evaporation. There should have 
overflowed from Step No. 1, during the 12-hour run, 1,872 c. c. 
of solution containing 1.5 percent of copper and 6.93 per.cent 
of H.,S 0 4 . When, as in Test No. 1, evaporation in No. 2 Step 
is compensated for by the addition of an equivalent amount of 
water, this becomes the step electrolyte with 0.375 percent of 
copper and 8.66 percent of H 2 SC) 4 . But in this test no water 
was added to No. 2 Step, so assuming evaporation to be the 
same in amount as during the preceding test, namely 925 c. c., 

2 


/ 


i8 


ROBERT RHEA GOODRICH. 


tliere should have overflowed to the sump 1,872 — 9 2 5 > or 947 
c. c. solution of the following composition: 

1872 

0.00375 X --— 0.00743 (or 0.743 percent) Copper. 


0.0866 X - = 0.171 (or 17.1 percent) H,S 0 4 . 

947 

Note that in Test No. 2, as well as in Test No. 4, the initial 
charge of electrolyte to all steps deviated slightly in H 2 S 0 4 com¬ 
position from the value which was calculated that it should have 
been. There was charged to No. 1 Step, a solution containing 
6.24 percent FL,S 0 4 , instead of 6.93 percent; to No. 2 Step and 
to Finishing Step, 15.4 percent H 2 S 0 4 , instead of 17.1 percent. 

The results of this test show that, when employing the Step 
Arrangement of Plant as outlined above and which differs in 
some details from the method used in Test No. 1, and when 
supplying initial charges of electrolyte of the compositions stated, 
these initial values remain constant in both copper and acid. The 
results of this test also show the performance when copper is 
deposited, using insoluble lead anodes. (See Photo No. 2; 
Table No. 6-B; Curve Sheet No. 6; Table and Diagram Example 
of Step Arrangement of Plant -A.) 


Test No. 3. 

But one electrolytic cell was used in this test. The intention 
was to introduce sulphur dioxide gas into the electrolyte during 
electrolysis in all subsequent work, so this test was run as a pre¬ 
liminary test to determine the best method of introducing the 
gas as well as the most suitable anode, in order to secure the 
best depolarizing effect. 


Kinds of Anodes. 

See Photo No. 6: Anodes Nos. 1, 2, 3, 4 and 5, top row, 
counting from left to right; Anodes No. 6, lower left corner; 
No. 7, lower right corner. 

Anode No. 1 was the lead Anode No. 6, used in Tests Nos. 
1 and 2, so modified that S 0 2 could be introduced into the elec¬ 
trolyte. Sufficient width and length was cut from one side and 




TREATMENT OF COPPER ORES. 


19 


the bottom of Anode No. 6, as formerly used, so as to permit 
of attaching a lead pipe, without increasing its size. This 

lead pipe was closed at the immersed end, and had its hori¬ 
zontal leg perforated on the upper side with three No. 
50 drill holes (about iVi n - diam.), equally spaced. It was 
thought that the placing of the perforation thus would give an 
even distribution of the gas over the surface of the anode. (See 
Photo No. 6, Anode No. 1 ; Table No-. 5-C.) 

Anode No. 3 was made of lead pipe to imitate in form the 
carbon anode (No. 7). The main part is a piece of a 34 -in. 
lead pipe. The bottom end was closed with a disc of lead in 
which was drilled a %- in. hole (the same size as the central 
channel in the carbon anode). A short piece of 34 ~in. lead pipe 
was forced into and burned to the upper end of the 34-in. lead 
pipe. To this pj-in. lead pipe rubber tubing was attached for 
supplying the sulphur dioxide gas. A soft rubber washer was 
placed on the pipe, near its lower extremity, to act as an insu¬ 
lator for preventing the occurrence of short circuits in the cell. 
Such a rubber washer was similarly placed on all other round 
anodes. This soft rubber washer had the same office as the 
hard rubber pegs of the flat lead anode (No. 6). The complete 
anode comprises a pair of these lead pipes which are rigidly 
held in a special copper clamp. (See Photo No. 6, Complete 
Anode No. 2; Table No. 5-C.) 

In Anode No. 3, a 34 ~in. lead pipe extends through a short 
section of a 34-in. ^ eac ^ pip e > which enables it to be held firmly 
in the standard copper clamp. A piece of charcoal free from 
flaws was cut exactly A m - square by 2^/2 in. long, and bored 
so as to fit snugly over the 34 -in. lead pipe. The sulphur dioxide 
gas enters the electrolyte from the bottom of the 34 “in. lead 
pipe. (See Photo No. 6, Anode No. 3; Table No. 5-C.) 

Anode No. 4 is of carbon. The carbon tube from which the 
anode was made is J^-in. outside diameter, A -in. inside diam¬ 
eter and 12 in. long. This is an electric arc carbon, incomplete 
in manufacture, and made by the National Carbon Co. Were 
it completed, the center channel would be filled with another 
grade of material, and then it would be a standard carbon for 
use in an electric arc for industrial purpose. This anode had 
a number of small holes made with a No. 50 drill, entering 


20 


ROBERT RHEA GOODRICH. 


radially to the center channel. (See Photo No. 6, Anode No. 4; 
Table No. 5-C.) 

Anode No. 5 differs from Anode No. 4 in that saw cuts were 
made instead of drill holes. These saw cuts extend half way 
through the carbon tube intersecting the center channel. (See 
Photo No. 6, Anode No. 5; Table No. 5-C.) 

Anode No. 7 differed only from Nos. 4 and 5 in that it had 
no lateral perforations, the gas entering the electrolyte from the 
bottom extremity of the center channel. (See Photo No. 6, 
Anode No. 7; Table No. 5-C.) 

The Depolarizing Effect of Sulphur Dioxide Gas. 

On the assumption that all chemical energy is transformed 
completely into electrical energy, it was found by calculation 
that when no depolarizer is used, in a copper sulphate electrolyte, 
the theoretical counter E. M. F. of polarization is 1.22 volt, and 
when sulphur dioxide is used as depolarizer, the theoretical 
counter E. M. F. of polarization is — 0.15 volt.* Thus theo¬ 
retically, sulphur dioxide gas should reduce the polarization by 
1.37 volt, changing the polarization of the cell from 1.22 volt 
opposing the generator to 0.15 volt, pulling with the generator. 
Consequently in this experiment a reduction of polarization of 
1.37 volt has been taken as the standard of perfection. In the 
calculation of the counter E. M. F. of polarization, the correc¬ 
tion term of absolute temperature times the temperature coeffi¬ 
cient of the electrical energy was neglected, because its value 
is unknown. The results obtained were somewhat influenced 
by the omission of this correction term. 

Polarization volts were determined as follows: During the 
test, when running at the current density desired, Reading 1 was 
taken of amperes and volts. By suitably manipulating the gen¬ 
erator field rheostat, together with the series resistance in the 
circuit, the current was reduced appreciably. New current value, 
together with corresponding volts, were quickly recorded, giving 
Reading 2. The current was quickly brought back to its run¬ 
ning value when Reading 3 was again taken of amperes and 
volts. Reading 3 should correspond to Reading 1. If it did not, 

* Trans. Amer. Electrochemical Society, Yol. 25, p. 230. 


TREATMENT OE copper ores. 


21 


the system was allowed to assume again its normal constant con¬ 
ditions and the same operation was repeated. All the readings 
were taken in a few seconds. 

Example illustrating determination of polarization: 

Reading i, at C. D. 8 amperes per scp ft., 

E = 2.637 amperes; V 1 = 0.710 volt. 

Reading 2, 

I 2 = 1.000 amperes; V 2 == 0.533 v °lt. 

E — I 2 = 1-637 amperes; V 1 — V 2 — 0.177 volt. 


Then 

V, =' E + E R 
V 2 = E + I 2 R 
V, — V 2 = (E — E) R 

. V, — Vo 0.177 

R = —- T " - = -——-= 0.108 ohm. 

ii —1 2 1.637 

E — Vi — ER — 0.710 — (2.637 X 0.108) — 0.425 volt. 

R equals ohmic resistance, made up of resistance in the solid 
conductor, contact resistance and resistance in the electrolyte. 
This is constant. 

E equals polarization voltage, constant at any one current 
density. It is assumed that during the time of taking Readings 
1, 2, 3, E does not vary. This is proven when Reading 3 checks 
Reading 1. Results were not recorded unless this was the case. 

On trying out the Anodes Nos. 1, 2, 3, 4, 5, 7, which were 
supplied with sulphur dioxide gas, it was found that with anodes 
Nos. 1, 2, 3 there was no reduction in polarization at all. The 
reason Anode No. 3 was tried was because, from a few ran¬ 
dom tests with the carbon anode, it had been learned that with 
this anode there would be beneficial reduction in polarization 
by the introduction of sulphur dioxide gas. It was thought 
that charcoal being quite porous might promote anode efficiency 
on account of its known property of occluding gases. A char¬ 
coal rod, when used as anode, was found to have too high 






22 


ROBERT RHEA GOODRICH. 


resistance to permit current to pass, therefore the combined 
lead-charcoal anode (No. 3) was made. This anode did permit 
current to pass, but, contrary to expectation, there was no bene¬ 
ficial depolarization. (See Table No. 5-C.) 

The carbon anodes Nos. 4 and 5, the former with radial holes 
and the latter with saw cuts, were tried with their submerged 
ends of central channel open. Then they were tried with their 
submerged ends plugged, forcing all the gas entering the elec¬ 
trolyte out through the radial holes and saw cuts. A variety of 
arrangements of holes and saw cuts were tried. (See Table 
No. 5-C.) Finally Anode No. 7, carbon anode with central 
channel extending through (not plugged) and with no other 
opening (neither radial drill holes nor saw cuts), was tried. 
The gas was delivered into the electrolyte entirely from the 
lower extremity of the anode. The depolarization efficiency 
of Anodes Nos. 4, 5, 7, were equally good compared one with 
the other. (See Table No. 5-C.) Therefore Anode No. 7, 
being the simplest in form as well as one of tl.ie most efficient, 
was selected to be used in Test No. 3 and in Test No. 4. 
Readings were then taken from which Curve Sheet No. 5 has 
been drawn. (See Tables Nos. 5-D, 5-E.) It is to be noted 
that during this test, sulphur dioxide gas was admitted through 
every anode, at such a rate that there was a gentle bubbling of 
the gas in the electrolyte. 

Depolarization was not always equally good with the same 
carbon anode. It was generally only with a new carbon anode, 
when on closing the circuit permitting current to pass at current 
density of 20 amperes per sq. ft., that a total voltage reading 
as low as 0.60 volt might be recorded. This extremely low 
voltage always rose to some higher value which remained con¬ 
stant during the test. A subsequent test with the same anode 
generally gave the first voltage reading higher, likewise, the 
reading of the constant value higher. When depositing copper 
with a current density of 20 amperes per sq. ft., at one stage 
of the experimenting it appeared that the total voltage could 
be held at 0.9 volt, at a later stage, 1.2 volts, while Test No. 4 
recorded 1.3 total volts. Apparently a new anode, when first 
put into use, is more efficient in reducing polarization than sub¬ 
sequently. This may partially account for variations in tests 


TREATMENT OF COPPER ORES. 


2 3 


on the same anode. From time to time there was considerable 
variation in the ohmic resistance of cell (due to contacts) which 
may account for some of this variation in total volts. 

finally, after completing all the regular readings for the 
curve sheet, a little further experimenting was done. Ten anodes 
were used in a test. Gas was admitted by the ten anodes, by 
four anodes only, and by six anodes only. The reduction in 
polarization was as good when admitting gas by four anodes 
or by six anodes as when admitting gas by all ten anodes. The 
only difference noticed was that the gas was required to flow 
through the anodes admitting gas somewhat faster than before 
—probably the volume of gas entering was no more. (See 
Table No. 5-C). 

In T est No. 3, when using different kinds of anodes, bene¬ 
ficial reduction of polarization was only secured when carbon 
anodes were used, indicating that the nature of the anode is 
important. When using the carbon anode, equally good results 
were secured with all forms of the carbon tube (National Car¬ 
bon Co. tube). It mattered not whether they were perforated 
by drill holes or saw cuts, for admitting and distributing the 
gas to the electrolyte, or whether it entered by every anode. All 
that was necessary was that the anode should be a carbon tube, 
with sufficient sulphur dioxide gas admitted to the electrolyte 
in some way. It appears that the quantity of sulphur dioxide 
gas required is that necessary to keep the electrolyte saturated 
with the gas. 


Solubility of Sulphur Dioxide in Water. 


H 2 0 Degrees C 

20 

30 

40 

50 

60 

70 

80 

90 

ICO 

SO 2 percent dissolved . . 

8.6 

7-4 

6.1 

4-9 

3-7 

2.6 

1-7 

0.9 

0.0 


H. O. Hofman, “General Metallurgy,” p. 880. 


Referring to the Tossizza Patent, which states: “I have thought 
to use .... insoluble anodes kept in contact with sulphur¬ 
ous acid and to utilize the known depolarization properties of 
the said sulphurous acid"; “These anodes can be made of car¬ 
bon.” From the above quotations, together with the results 



























24 


ROBERT RHEA GOODRICH. 


obtained in this test, I infer that Tossizza experimented only 
with carbon anodes. (See Photos Nos. 4, 5, 6; Tables Nos. 
5 ~A, 5 - b > 5-C» 5 _B) ’ 5-H; Curve Sheet No. 5.) 

Test No. 4. 

The operation of this test differed from Test No. 2 only in 
that carbon tube anodes were used in place of the lead anodes, 
and sulphur dioxide gas was introduced into the electrolyte. 
Were depolarization perfect in operation, then for each equiva¬ 
lent of copper deposited, two equivalents of H 2 S 0 4 would appear 
instead of one when sulphur dioxide gas is not used. (See 
“Trans. Amer. Electrochemical Society/’ Yol. 25, p. 230.) In 
this test, the initial charges of electrolyte to the different steps 
had the degree of acidity which had been determined suitable 
for Tests Nos. 1 and 2 where S 0 2 gas was not employed. Con¬ 
sequently, when using sulphur dioxide gas as depolarizer, the 
electrolytes became more acid in proportion to the anode efficiency 
as the electrolysis progressed. The tabulation of Test No. 4 
shows a decided increase in acidity. (See Table No. 6-B.) 
When the anode efficiency is known and when the initial elec¬ 
trolyte charged to the cells is made correspondingly more acid, 
then the electrolyte should remain constant in acid contents. 
Anode efficiency is discussed fully in Test No. 3, and compu¬ 
tations are explained in Tables Nos. 5-D, 5-E. It was there 
estimated by percentage reduction in polarization rather than 
by increased acidity, as with a fine electrical instrument this 
method was thought to be more accurate than analytical meth¬ 
ods. By the electrical method, results were recorded instanta¬ 
neously, while the determination of the sulphuric acid by chem¬ 
ical analysis would require days to secure a series of results. 

Referimg to the tabulation of lest No. 4* where the average 
lesults of leadings taken half-hourly are recorded, quite a saving 
in power in Nos. 1 and 2 Steps of Test No. 4 over Test No. 2 
will be seen. 

Finishing Step. 

1 he curve sheet of the Finishing Step represents the per¬ 
formance of this step of the process in Tests Nos. 2 and 4. 


TREATMENT OF COPPER ORES. 


25 


The 5,210 c. c. of electrolyte, containing 0.743 percent copper 
and 15.4 percent H 2 SQ 4 , with .current of 1.648 amperes passing, 
required theoretically, 19.85 hours to completely deposit the 
total copper contents. In these tests, when carried to the 19.85 
hour-point, the results were: 

Test No. 2 Test No. 4 

Copper deposited, percent of that sup¬ 


plied to Finishing Step . 93.5 90.2 

Average current efficiency . 93.5 90.2 


While the extraction is high and current efficiency good in 
both tests, it would not be desirable to carry the electrolysis to 
to this point (19.85 hour-point) in Test No. 2, because the last 
of the copper comes down spongy and falls to the bottom of 
the cell. In Test No. 4, the electrolysis may be carried to the 
19.85 hour-point and a good cathode produced on account of the 
presence of S 0 2 in the electrolyte. But even in Test No. 4, 
this point is the limit of feasible electrolysis. At the 20 hour- 
point, great blotches of sulphide of copper formed on the 
cathode. Test No. 2 may be carried to the 18th hour and pro¬ 
duce a good, firm deposit of copper. In Test No. 2, when car¬ 
ried to the 18 hour-point, the results were: 

Test No. 2. 

Copper deposited, percent of that supplied to 


Finishing Step . 85.5 

Average current efficiency. 94-4 


Note that there is a decided saving in power by the use of 
sulphur dioxide gas with the carbon anodes, in the Finishing 
Step of Test No. 4. In Test No. 4, there is no rise of polari¬ 
zation as the copper contents becomes depleted, while the total 
voltage becomes 2.2 volts in Test No. 2. 

The middle row of Photograph No. 6 consists of six cathodes. 
Counting from left to right, the first group of three (Nos. 1, 2, 
3) was used in Tests Nos. 1 and 2; the second group of three 
(Nos. 4, 5, 6) was used in Test No. 4. Immediately below 
Cathode No. 2, is one of the lead anodes used in Tests Nos. 
1 and 2; immediately below Cathode No. 5, is one of the carbon 
anodes used in Test No. 4. Cathodes Nos. 1 and 4 are from 
No. 1 Step of their respective tests; Cathodes Nos. 2 and 5 are 






26 


ROBERT RHEA GOODRICH. 


from No. 2 Step of their respective tests; and Cathodes Nos. 

3 and 6 are from the Finishing Step of their respective tests. 

The performance in Test No. 4, in pounds of copper deposited 

per kw. hour—between 1.82 and 3.7 lb.—is not equal to that 
in Test No. 3. Owing to the construction, electrodes light in 
weight, and resting on bus-bars, the contact resistance was ab¬ 
normally high. Special precaution was taken in Test No. 3 to 
largely eliminate the contact resistance. Even in Test No. 3, 
the ohmic resistance is to that of practice as 0.108 is to 0.062. 
Consequently it is believed that even the showing of Test No. 3 
can be bettered in practice, and with this plant on repetition. 

Test No. 1 showed the performance of lead anodes when used 
with the step system. Test No. 2 was a modification of Test 
No. 1, seeking to place the operation on a more practical basis. 
Test No. 3 compared the performance of different kinds of in¬ 
soluble anodes with and without sulphur dioxide gas. Test No. 

4 showed the performance of carbon tube anodes with sulphur 
dioxide gas introduced into the electrolyte, when used with the 
step system. The operation of the Finishing Step with S 0 2 
gas showed a remarkable saving of power as well as a better 
deposit of copper over that of the Finishing Step in Test No. 
2 with lead anodes and no gas. (See Photos Nos. 2, 6; Tables 
Nos. 6-A, 6-B, 6-C ; Curve Sheet No. 6.) 

Step System of Electrolytic Precipitation. 

1 he copper solution used in the preceding experiments was 
obtained by dissolving copper sulphate in water. The electro¬ 
lytic plant was not supplied with copper solution obtained by 
leaching the ore, because the amount of ore roasted was insuffi¬ 
cient to keep the electrolytic plant in operation for the length 
of time desired. Moreover, it is believed that the only addi¬ 
tional information that could have been obtained by treating 
copper solution resulting from ore leaching, would have been 
the effect of accumulation of impurities in the electrolyte. In 
order to determine the effect of accumulation of impurities in 
the electrolyte, and to inaugurate the necessary purification meth¬ 
ods, such as chemically purifying the electrolyte or wasting a 
sufficient amount of barren solution, a greater amount of mate¬ 
rial than was at hand would have been required, as well a« a 


TREATMENT OF COPPER ORES. 


2 7 


long campaign of operation. Much information is available re¬ 
garding the maintaining of the purity of solutions in a cyclic 
process. The analysis of the ore together with the results of 
the leaching tests, and the acid consumption, enables one to judge 
the quantity of impurities entering the leaching solution in the 
treatment of the ore now under discussion. 

I he electrolysis of the copper sulphate solution was conducted 
in steps. All the electrolytic cells were connected in series, 
while the electrodes of the individual cells were connected in 
parallel. Each step maintained constant the composition of the 
electrolyte—both in copper and acid. 

The cells comprising each step have two circulations of the 
electrolyte. There is one circulation, which is designated the 
step circulation, in which the step circulating pump draws the 
electrolyte from the last cell of the step sending the electrolyte 
to a more elevated tank (temperature tank) in which it may be 
heated (maintaining the temperature of the system constant), 
and from which it overflows and gravitates into the first cell of 
the step. The electrolyte then passes on through the series of 
cells in the step and finally again to the same circulating pump. 

There is the other circulation, which is ordinary progressive 
movement of the solution through the plant. The copper solu¬ 
tion resulting from the leaching of the ore, or otherwise ob¬ 
tained, is run, together with the discharge of the No. i Step 
circulating pump, into the elevated temperature tank of the 
No. i Step. Thus more solution enters the first cell of the step 
than is drawn away by the circulating pump, consequently an 
equivalent amount of solution must leave the last cell of the 
step by way of the overflow discharge. This overflow dis¬ 
charge passes on and joins the discharge of the circulating pump 
of No. 2 Step, and enters the elevated temperature. tank con¬ 
nected with that step. Finally the last cell of the step pre¬ 
ceding the Finishing Step, overflows an amount of solution equal 
in amount to the inflowing copper solution fed to the elevated 
temperature tank of the No. i Step. 

The copper contents of the solutions is maintained in geo¬ 
metrical ratio or in arithmetical difference from that of the copper 
solution fed to No. i Step to that of the electrolyte of the step 
preceding the Finishing Step. 


2 8 


ROBERT RHEA GOODRICH. 


The experimental plant was designed so that a variety of fac¬ 
tors could be used. (See Tables and Diagrams Examples of 
Step Arrangement of Plant.) Let us select for a plant a solu¬ 
tion resulting from the leaching of the ore with copper contents 
o'f 6.0 percent and the geometrical ratio 1/4 for the copper con¬ 
tents of the electrolyte of the different steps. (See Table and 
Diagram Example of Step Arrangement of Plant -A.) Then 
the electrolyte of No. 1 Step will contain 6.0/4 equals 1.5 per¬ 
cent copper, and 6.93 percent H 2 S 0 4 . Let us further assume 
for No. 1 Step twelve cells in series, each with one unit of area 
of cathode surface. Next select the most suitable current den¬ 
sity for this step. The copper contents of the electrolyte is the 
governing factor in making this selection, since it is desirable 
always to maintain current density proportional to copper con¬ 
tents of the electrolyte. The selection of current density fixes 
the strength of current. The copper solution is fed to this step 
at such a rate that the copper deposited on the cathode amounts 
to three-quarters of the entering copper. The electrolyte cir¬ 
culating in the step thus remains constant in copper contents 
and in acid, and the solution overflows in volume equal to that 
of the inflowing feed solution. 

The solution fed to No. 2 Step, being the overflow of No. 1 
Step, contains 1.5 percent copper and 6.93 percent H 2 S 0 4 . 
Since the electrolyte of this step contains one-quarter as much 
copper as that of the preceding step, then in order to maintain 
the electrolyte of this step constant one-quarter as much copper 
must be deposited on the cathodes of this step. This is accom¬ 
plished by placing one-quarter as many cells in series, namely, 
using three cells. These cells should each have four units of 
cathode area, that is, a cathode area four times as great as that 
employed in each cell of No. 1 Step, since in this Step (No. 2) 
the electrolyte contains one-fourth as much copper, which re¬ 
quires a current density one-fourth that of No. 1 Step to be 
employed. 

It is seen that the solution overflowing from No. 2 Step con¬ 
tains but 6.25 percent of the original copper contents. This 
solution in which the acid has been regenerated may pass on 
and be used to leach ore. That it contains a small amount of 


TREATMENT OF COPPER ORES. 


29 


copper is no detriment since this copper will return in the en¬ 
riched solution and will not be lost. 

In case it should be desired to extract the copper to the last 
trace, one more step—a finishing step-^-may be added to the 
plant. This step would be supplied intermittently with a charge 
of solution from the sump tank, where the solution overflowing 
from No. 2 Step has collected. This charge of solution would 
be circulated in the Finishing Step by a circulating pump in the 
same manner as in the preceding step but with neither feed nor 
overflow, until all of the copper contents is deposited on the 
cathode. The solution, barren in copper, would, be withdrawn 
from the system, after which a new charge would be supplied. 

The number of cells in series in the Finishing Step would 
be one-third the number used in No. 2 Step. Then this step 
must operate continually to deposit all the copper sent to it by 
the overflow of No. 2 Step, since the amount of copper con¬ 
tained in the overflow solution of No. 2 Step is one-third the 
amount deposited in that step. The cells of the Finishing Step 
should each have (in this example there is but one cell) eight 
units of cathode area, that is, twice the cathode area and one- 
half the current density as that employed in each cell of No. 2 
Step, as in this step (Finishing Step) the electrolyte has on an 
average one-half the copper strength. 

Even when the Finishing Step is employed, it may be desi¬ 
rable to stop the electrolysis somewhat short of complete ex¬ 
traction. (This matter is discussed in connection with Tests 
Nos. 2 and 4.) 

In order to obtain the figures given above, it is necessary 
to compensate for evaporation, which at the temperature em¬ 
ployed (50° C.) and with the large surface of solution exposed 
is considerable. In Test No. 1, evaporation was compensated 
for by the addition of water to the temperature tank of each 
step, equal in amount to that of the evaporation. Thereby was 
demonstrated the feasibility of maintaining the composition of 
the electrolyte constant in composition, both in copper and acid, 
during the electrolysis, and with copper contents of the electro¬ 
lyte of the different steps in the above described ratios (4:1). 

In Tests Nos. 2 and 4 the plan of operation was slightly 
modified from that followed out in Test No. 1 in order to some- 


30 


ROBERT RHEA GOODRICH. 


what simplify the operations. To the temperature tank of No. 
i Step was fed a more dilute copper solution, which amounted 
to the original volume of the 6.0 percent copper solution plus 
the required amount of water to compensate for evaporation 
of No. 1 Step. Consequently No. 1 Step operated in every way 
exactly as it did in Test No. 1, thus maintaining the compo¬ 
sition of its electrolyte the same and causing an overflow equal 
in amount and composition to that in Test No. 1. This calling 
for a more dilute copper feed solution might be an advantage 
in practice, as it might be easier to secure than the one of 
higher copper contents. Further dilution of the copper solution 
fed to No. 1 Step, to compensate for evaporation in No. 2 Step 
is not permissible as it would derange the constant conditions 
desirable to be maintained in that step. Moreover for simplicity 
of operation, water was not added to No. 2 Step to compensate 
for evaporation, consequently the copper contents of the elec¬ 
trolyte of No. 2 Step differed from the original geometrical ratio. 
A higher value of both the copper and the acid contents ob¬ 
tained. Such higher values of both copper and acid, however, 
remained constant. The higher copper value would permit 
of a somewhat smaller electrode surface and higher current den¬ 
sity per cell than was called for in Test No. 1. 

The ideal of the copper hydro-metallurgist is to secure, in 
the hydro-electrolytic extraction of copper from its ores with 
insoluble anodes, conditions comparable with those which obtain 
in the electrolytic refining of copper with soluble anode. By 
means of the step system it is believed that these conditions are 
more nearly approached than they have been heretofore. The 
desirable conditions are: a—Constant composition of electrolyte 
with current density adjusted to suit composition; b—Small 
power consumption. 

The step system arrangement of the electrolytic cells accom¬ 
plishes “a" and also permits the electrolyte to be circulated at 
any rate desired as in electrolytic refining of copper. A rapid 
rate of circulation in the individual cells which is made possible 
by the step system, together with the adjusting of the current 
density proportionate to the copper contents (which is held con¬ 
stant in each step) makes possible the securing of high current 
efficiency. 


TREATMENT OF COPPER ORES. 


31 


As most copper ores contain sulphides', they should be roasted 
prior to leaching. Sulphur dioxide gas is evolved. I11 such 
cases it may be desirable to pass this gas through the electro¬ 
lytic cells, utilizing it as a depolarizer and at the same time pro¬ 
ducing additional sulphuric acid for the leaching. Ordinarily 
in leaching copper ores, the solution takes up some iron from 
the ore. Were no sulphur dioxide gas introduced into the cell, 
the ferrous sulphate would become oxidized to ferric sulphate 
at the anode. The ferric sulphate formed would be carried by 
circulation of the electrolyte to the cathode where it would dis¬ 
solve some of the deposited copper, again becoming ferrous 
sulphate, after which the cycle would be repeated. Current 
efficiency would thus be decreased. But the sulphur dioxide gas 
when employed with a suitable insoluble anode, besides main¬ 
taining a high current efficiency, is beneficial in reducing the 
power consumption to a point more comparable with that con¬ 
sumed in the electrolytic refining of copper with soluble anode. 
(See Curve Sheets Nos. 5, 6.) 

Increased areas of electrodes to give reduced current density 
for corresponding depletion in copper contents of the solution 
in the process of electrolysis have been used heretofore. In these 
processes, however, there is but one circulation of the electrolyte, 
namely, that through the plant progressively ; the solution be¬ 
comes depleted of the copper during its passage. Such a cir¬ 
culation would necessarily be slow and insufficient for securing 
the best results. Moreover the electrolyte, due to the slow prog¬ 
ress through the cells, would vary in composition in different 
parts. So, although in the ordinary processes the aim is to main¬ 
tain the current density proportional to the copper contents, it 
has not really been accomplished. 

The step arrangement of the plant is such that— 

1. Rapid circulation is maintained by the step circulating 
system. 

2. The composition of electrolyte remains absolutely constant 
in each step of the plant. 

3. The current density is held strictly proportional to the 
copper contents of the electrolyte. (See Tables Nos. 5 _ A, 5~E, 
5-C, 5-D, 5-E, 6-A, 6-B, 6-C; Curve Sheets Nos. 5, 6; Tables 
and Diagrams Examples of Step Arrangement of Plant.) 


3 2 


ROBERT RHEA GOODRICH. 


SUMMARY. 

It was desired to select an Arizona problem, so the treatment 
of a porphyry copper ore seemed to be one of the most im¬ 
portant. This kind of ore when treated for the extraction of 
its copper by mechanical concentration and smelting, the methods 
most generally employed, yields 66 percent or less of its copper. 

In these experiments, therefore, the aim was to determine 
a better method of treatment. T he method selected for investi¬ 
gation was one in which metallurgists are at present doing much 
experimenting in the hope of demonstrating the superiority of 
leaching and electrolytic precipitation over earlier practice, 
namely: i, Oxidizing roast; 2, Leaching with dilute sulphuric 
acid; 3, Electrolytic precipitation of the dissolved copper. 

While it was planned to carry on a complete systematic test 
of the ore, thereby enabling one to determine the suitability of 
this ore to the hydro-electrolytic treatment, special stress was 
laid on roasting and on electrolytic precipitation using sulphur 
dioxide to lessen the consumption of power and to produce addi¬ 
tional sulphuric acid. 

It has been demonstrated in these experiments that it is feasible 
to roast successfully the ore, so that when leached with hot dilute 
sulphuric acid, the entire copper contents may be obtained in 
solution. 

Some of the practical engineers say that they secure but little 
beneficial reduction in power consumption when using sulphur 
dioxide gas as a depolarizer. Tossizza says, “I have thought 
to use . . . insoluble anodes kept in contact with sulphur¬ 
ous acid, and thus to utilize the known depolarization properties 
of the said sulphurous acid. These anodes can he made of car¬ 
bon, and in this case the sulphurous acid can be introduced 
outside the anode or in the interior thereof“One can thus 
obtain a very beautiful deposit of pure electrolytic copper . 
with a sufficient intensity at a volttage of about six-tenths of 
a volt." The experiments reported under Tests Nos. 3 and 4 
corroborate Tossizza’s statement in that the method of intro¬ 
ducing the sulphur dioxide into the electrolyte is unimportant, 
provided only that it be introduced in sufficient quantity to keep 
the electrolyte saturated. 


TREATMENT OF COPPER ORES. 


33 


Regarding the saving of power (depolarization), these experi¬ 
ments show depolarization by the use of sulphur dioxide only 
when used in connection with carbon anodes, while Tossizza, 
although specifically suggesting a carbon anode, intimates that 
the same may be obtained with other insoluble anodes. Tos¬ 
sizza does not state the current density at which he was operating 
when depositing copper with the extremely low voltage of six- 
tenths of a volt, although “with a sufficient intensity” might 
mean commercial current density. Experiment Test No. 3 shows 
likewise that copper deposits continuously at six-tenths of a 
volt when current density equals 5.8 amperes per square foot, 
depositing 4 pounds of copper per kw. hour. (See Curve Sheet 
No. 5.) It is to be noted that although beneficial depolarization 
and consequent saving of power is secured at all current den¬ 
sities, when SO 2 is introduced into the electrolyte, it is rela¬ 
tively not the same in amount but decreases as the current den¬ 
sity increases. 

It is to be hoped that future experiments will demonstrate 
how to secure the same beneficial depolarization with sulphur 
dioxide when electrolyzing with high current densities. Like¬ 
wise whether if possible and how beneficial depolarization by 
sulphur dioxide may be secured with other kinds and types of 
insoluble anodes. 

When it is desired to extract the copper from the electrolyte 
down to a small trace, sulphur dioxide is very beneficial, not 
only in reducing the power consumption but in causing the 
copper to deposit more firmly on the cathode, which when sul¬ 
phur dioxide gas is not used forms as a spongy deposit toward 
the last of the electrolysis. On attempting to extract the last 
trace of copper from the electrolyte, sulphide of copper forms 
on the cathode. 

Regarding the lead anodes, there is no depolarization when 
used with or without sulphur dioxide gas. There was no loss in 
their weight during the test. 

When starting out on this line of investigation, it was found 
necessary to design and construct a complete plant on a minia¬ 
ture scale. In doing this, a novel idea appeared—the step ar¬ 
rangement—which greatly facilitated the work in all stages of 
the process, especially in the finishing step where the copper 


3 


34 


ROBERT RHEA GOODRICH. 


was extracted down to a trace with the production of a good 
firm cathode, when sulphur dioxide gas was used. 

Test No. i demonstrates, when using the step system, the 
feasibility of depositing copper continuously from the electro¬ 
lyte, while at the same time the copper contents of said elec¬ 
trolyte remains undiminished. A plant composed of several 
steps of such cells may be operated so that a liquor strong in 
copper flows into the upper step of the plant continuously from 
the leaching vats at such a rate that a liquor depleted of its 
copper, in which the acid solvent has been regenerated, outflows 
from the plant continuously to the leaching plant. The advan¬ 
tages of a step arrangement of plant are apparent in that the 
electrolysis is conducted under constant invariable conditions in 
each cell of the respective steps, as well as in every part of the 
cell. The cathode area of a step may easily be adjusted so that 
the most desirable current density is secured (as determined 
by experience in operating). It is to be noted in step arrange¬ 
ment of plant, the electrolyte may be circulated in any of the 
steps at any desired rate. This has been shown to be beneficial 
in promoting high current efficiency. 

The experiments demonstrate that the porphyry ore, when 
treated by a hydroelectrolyic process, will yield its entire copper 
contents as a good grade of cathode copper. It is therefore 
hoped that by using a hydroelectrolytic method, lower grade cop¬ 
per porphyry .deposits may be worked than formerly could be 
worked by methods of concentration followed by smelting. 


Epitome. 

i. The most suitable temperature for roasting the Arizona 
porphyry copper ore, containing sulphides, in order to render 
it amenable to acid leaching methods, is between 6oo° C. and 
725 0 C. The more finely ground the material the shorter the 
time required for the roasting so as to produce the maximum 
amount of soluble copper: materials which will pass through a 
20-mesh screen and remain on 80-mesh require about two hours’ 
roasting at 6oo° C. to 725 0 C.; and when ground to pass through 
an 80-mesh screen the time required is about 1% hours. If the 
roasting is concluded at temperatures above 8oo° C., the oxidized 


TREATMENT OF COFFER ORES. 


35 


copper is converted into a compound which is insoluble in dilute 
sulphuric acid. The longer the roasting is conducted above 
800° C., the greater the amount of insoluble copper produced. 

2. A heated solution is necessary to leach efficiently the cop¬ 
per from the roasted material. A 10 percent H 2 S 0 4 solution at 
ioo° C. leached out, in from 3 to 6 hours, all the copper from 
all roasted materials (through 40 on 80-mesh, through 80-mesh, 
whole through 20-mesh), except material through 20 on 40-mesh, 
in which case the extraction was not so high. 

3. The nature of the anode is an important factor in securing 
depolarization by sulphur dioxide gas. 

4. Depolarization with consequent saving in power is accom¬ 
plished when using sulphur dioxide gas with a carbon anode, 
while there is no depolarization with a lead anode. 

5. The depolarization by sulphur dioxide gas, even with car¬ 
bon anodes, does not reach the theoretical amount, being between 
45 and 65 percent. 

6. The amount of depolarization effected by sulphur dioxide 
gas when used with carbon anodes varies with the current den¬ 
sity, being a maximum at low current density. 

7. The method of introducing the sulphur dioxide gas into 
the cell is unimportant. All that is necessary is that it be intro¬ 
duced in some way, and in such quantity that the electrolyte is 
saturated with the gas, so that there is some escaping by bubbling 
at the surface. 

8. A smoother deposit of copper forms when using sulphur 
dioxide gas as a depolarizer than when not using it. 

9. In the electrolysis of an acid solution of copper sulphate, 
when sulphur dioxide is not supplied to the cell and when one 
is endeavoring to carry electrolysis to the point of complete ex¬ 
traction, a soft spongy deposit begins to form on the cathode 
before the complete extraction of the copper is effected, ffhere 
is also a considerable rise in polarization when the copper con¬ 
tents of the electrolyte becomes low. 

10. In the electrolysis of an acid solution of copper sulphate, 
when sulphur dioxide is supplied to the cell, the copper con¬ 
tents of the electrolyte may be reduced to a very small trace with 





36 


ROBERT RHEA GOODRICH. 


the formation of a good, firm cathode, without rise in polariza¬ 
tion toward the end. Current density and energy efficiency re¬ 
main high to the end. It is only when prolonging the operation 
beyond the time when but a small trace of copper remains that 
sulphide of copper forms as a thin coating on the cathode. 

11. Lead anodes do not peroxidize or deteriorate appreciably 
when used with or without the introduction of sulphur dioxide 
into the electrolyte. 

12. A novel idea—step arrangement of process—makes fea¬ 
sible the depositing of copper continuously, while the copper 
contents of said electrolyte in the respective steps remains con¬ 
stant. A plant composed of several steps of such cells may be 
operated so that a liquor strong in copper flows to the plant con¬ 
tinuously, and the liquor which outflows from the plant is de¬ 
pleted of its copper. 

13. The circulation of the electrolyte by the step arrangement 
increases the current efficiency. Such rapid circulation is not 
possible in plants as ordinarily arranged. 






TREATMENT oE COPPER ORES 


37 





Photo-micrograph Gi. 
Magnification 30 diameters. 








38 


ROBERT RHEA GOODRICH 



Photo-micrograph G 2 . 

Magnification 30 diameters. 






TREATMENT OF COPPER ORES 


39 





Photo-microc.raph G 3 XN. 

Magnification 30 diameters. 






Photograph No. 



testing table. 



Photograph No. 


TREATMENT OF COPPER ORES 


41 



ELECTROLYTIC cells. 










42 


ROBERT RHEA GOODRICH. 


Photograph No. 3. 



FRAME. 




























treatment of copper ores 


43 



CELL ASSEMBLED. 













44 


ROBERT RHEA GOODRICH. 


* 


Photograph No. 5. 




I 

CELL TAKEN APART. 













Photograph No. 6 . 







ELECTRODES. 






46 


ROBERT RHEA GOODRICH. 


Tab l c No. / 


Screen Analysis of Crashed Ore 


'C 

VJ 

s: 

"N* 

V. 

* 

V 


£ 


k) 

$ 


r~ 

V. 

V& 


<k 

£ 

vj 

v. 

Cl 


$ S 
a 

h C 

A) 

*V ^ 

'k - 

y * & 

^ VJ 

^ <u ^ 

^ u ^ 

- i $ 

-5 ^ 

<U 


s *> 

5 "5 <o 


o 


^ «+* 

VJ & 


k 


on 20 


thru 20 on 3o 


3 6 3 


28. 


23 . 


30 .. 4-0 


2 ZG 


/ 7.4 


45. A 


40 .. SO 


/ 4-9 


II. 5 


SG.9 


•i S o ,i Go 


2 G 


2.0 


S3. 9 


60 


70 


0.7 


S9.G 


7o 


So 


113 


8-7 


63.3 


.. SO -, 9o 


24- 


1.3 


70 . / 


_9o ,. / OO 


d 7 


4-3 


74.4 


.. loo .. /So 


/ 0 3 


3. O 


82.4 


thru /So 


2 26 


total 


/ 2 96 


/ 7 - & 


/ o o. 


/ O o. 




































treatment oe copper ores. 


47 


Table No. 2_A 


ft o as i 

fm S tn the Obs Muffle Furnace a1 temperature 500 ° C. 

i 

r 

1 

■— 

\ 

c 

<U 

<U 

N 

s 

* 

C Si 

* T 

< r 
i 5 

$ ^ 

R <S 

R 

A 

V 

o) A 

R ^ 

<5 

K '-v 

C 

U 

VJ 

C 

<0 

Q. 

1 

R> ^ 

R 

s 

X 

V. 

-£ -5! 

5 A 

5s 

,A 

^ v 

5 c 

v 

4? £ 
t A 
o c 1 
- 4e 

3 8 

■> 

O 

v. 

X 

$ 

A o) 
A \ 

a e 

$ $ 

■u 

5 5 

« £ 

5 * 

5 A 
C) >J 

A-. 

o 

N 

X 

n 

A s 

iS <3 

£ V 

« 0 
* • V 

4! * 

-0 c 

3 % 

\j 

H <» 
^ A 
£ 1 
iS3 

A"' 

O 

X 

3 „ 

4» 9 . 
-5 5 

<3 ^ 

* 

R O 

j. *W 

A 

3 

-! e 

<1 il 

3 'J 
\ V. 
0 01 

« *■ 

A-i 

0 

X 

3 ■> 

^ * 
* 3 

•v. ^ 
Q V. 

A +*• 

At 

Sc 
^ t 

A 

0 

X 

|-a 

•*» 0 
<3 5 

3 3 

'k! 

Ai, 

1 

R 

R 

i? 

^ Oj 

A A 

Sfarv Ore 

thru 2o on 40 

O. 

/4.IO 

0. 

1.65 

7.65 

4.37 

6.02 

27.4 

ffatu Ore 

thru 8 o 

0. 

9.40 

0. 

!.5o 

/.5o 

4/8 

5.68 

26.5 

/ _ a 

thru Zo on 40 

0.5 

3.3 6 

0.41 

4.02 

4.43 

196 

6.41 

69X 

/ _ b 

thru 8o 

0.5 

r.o4 

/.OZ 

4.66 

5-70 

0.41 

6.11 

933 

Z- a 

thru ZO on 4-0 

1.0 

/■66 

1.87 

4.31 

6.18 

0.61 

6.79 

9‘. i 

Z _ b 

thru 8o 

1.0 

1.30 

1.87 

3.44 

5.31 

0.7s 

6.06 

87-8 

3_ * 

thru Zo on 4o 

1.5 


2.5o 

3.88 

6.38 

0.46 

6.84 

933 

-3 _ b 

thru 8o 

/.5 


2.0 £ 

3.28 

5.34 

O .70 

6 .04 

38.4 

4- a 

thru 20 on 4o 

2.0 

179 

2.5o 

3.62 

6.12 

0.65 

6. 77 

90.5 

b 

thru 8o 

2.0 

1.4 6 

/ 92 

3. IS 

5.07 

0.95 

6.02 

84 .3 

5 _ * 

thru 20 on 40 

2.5 


2.52 

3 59 

6.11 

0.15 

6.2 6 

977 

£- b 

thru So 

2.5 


2.01 

3.59 

5.60 

0.46 

6.06 

92.4 

6 _ a 

thru Zo on 40 

3.0 

1.65 

2 SO 

3.64 

6.14 

027 

6.41 

95Q 

6_ b 

thru 8o 

3.0 

/ 43 

/■99 

3.37 

5.36 

0.68 

6.o4 

888 

















































48 


ROBERT RHEA GOODRICH 


Table No. Z _ f? 


Ffoasfmj m the 6as 

Muffle. Furnace a 

t temperature 

600 

°C. 

o' 

N 

Q. 

$ 

* 

I 

X 

3 

x» 

c 

Qj 

V 

<$ 

0 « 

V 

1 

N 

X. 

\ 
c £ 

5s * 
o ^ 

^ V 

1 0 

$ 

c Oi 

<J c: 

V <0 

11 

K 

-k 

<U 

V) 

C 

QJ 

1 

'-i 

O 

X 

X 

£ ^ 

§ <*» 

x> <• 

V) 

* *! 
■s H- 

b 

a i 

r "<r' 1 

<3 

X 

X 

X 

3; W 

S k 
* 5 

5 <5 

X ^ 

o ^ 
*; 

c: * 

X \J 

VJ 

« Si 

-5 v3 
o ^ 

3 ^ 
k> ^ 

'"o '"' 1 

<3 

X 

X 

3 « 
^ \ 

s o 

X ^ 

§ ° 

-0 c: 

3 <6 

x Xl 

0 k 

s o 

x 'k 
£ 1 

o '" 1 

<3 

X 

X 

.3 

\3 4) 

4$ 1 

Xl 5 

-5 3 

X X 

0 O 

X 

i i 

3 

so 

^ c 

-0 Qj 

3 Vi 

x X 

0 3j 

S ^ 

X 

o 

— 

X 

3 -x 

^ S' 
5 5 

0 X 

X Q 

X- X 

1 < 

3 c 
Vi v 
^ u 

-X * 

0 ^ 
K 

o 

X 

X 

>5 „ 
u -c 

X a. 

si c 

X 3 
x 

H 

OJ 5 

■u 

1 

s: 

0 

- c 

5 5 

J V 

X (jj 

5 * 

Ffurv Ore 

thru 20 on 40 

O. 

14.10 

O. 

1.65 

1.65 

4.37 

(S.oz 

274- 

ffarr Ore 

thru 30 

0. 

9.40 

O. 

1.50 

/.5o 

4. Id 

5.68 

26.5 

1 -a 

thru 20 on 40 

O.S 

4.68 

0.22. 

3.30 

3.52 

2 66 

6.18 

5J.O 

7 - t> 

thru do 

0.5 

3.10 

0.73 

3.86 

4.59 

1.59 

6.13 

74.4 

Q - a 

thru 20 on 4o 

1.0 

0.15 

0.07 

6.58 

6.65 

0.0 5 

6.7o 

99.4 

Q - b 

thru So 

1.0 

0.13 

0.0 7 

5.61 

5.68 

0.48 

6.16 

92.2 

9 _ <7 

thru 20 on 4o 

t.5 


0.07 

6.49 

656 

O.l 1 

6.67 

99-4. 

9- b 

thru do 

f.5 


O.IO 

5.75 

5.85 

0.33 

6.13 

94.7 

t o _ a 

thru 2o on 4o 

2.0 

0.03 

o. to 

6.53 

6.63 

0. 12 

6.75 

98.3 

to _ b 

thru So 

2.0 

0.16 

O.IO 

5.66 

5.76 

0.42 

6.13 

93.3 

/ / _ a 

thru 20 on 40 

2.5 


O.IO 

6.67 

6.77 

0. 

6.77 

100.0 

II - h 

thru 60 

2.5 


0.12 

5.63 

5.75 

0.41 

6.16 

93.3 

IZ-a 

thru 2 0 on 4o 

3. o 

o. 05 

0.12 

6.48 

6.6o 

O.l 7 

6.77 

97 S 

12- b 

thru So 

3.0 

O.l / 

0. !Z 

5.66 

5-80 

0.36 

6.16 

94.2 





































TREATMENT OF COPPER ORES. 


49 


Ta b / e No. 2 — C 


Ffoastiny mthe Gas Muffle Furnace at temperature 725°C. 

o 

<u 
—. 

\ 

\ 

>*» 

V 

HJ 

-K 

Qj 

N 

s 

"N. 

* 

V. 

k <0 

* 4 

o 3 
< v. 

1 S 

$•0 

5: 

■n -v. 

1) 

R: 

<« 

.s <3 

K ^ 

■v. 

Oj 

u 

V 

<u 

<E 

1 

'n 

<3 

N 

X 

k 

§ Ck 

J $ 

5 <3 

c- ^ 
5 $ 
4) ^ 

H 

\ i 1 

Vo i 

o 

v. 

X 

$ 

<u 

s \ 
* 5 
$ * 

Q u 

•u 

c k 

■sS 

0 ^ 

3 'S 
VJ >s 

r 'cT'' 1 

O 

X 

3 <u 

^ S' 

Q § 

e “ 
<• 

3 

3 <k 
^ vj 

S 

v. <\ 

$ ' 
.0 3 
k Vj 

r o-. 

o 

X 

a „ 

4! 9. 

•3 * 

0 ** 

1 

3 

VJ 

-S t 

<3 (0 

3 * 

0 y 

r 'o'' 1 

X 

3 

Vo <3. 

<a v 

0 

"k* Vr 

1 ^ 
3 r 

V) J 
- 

•k *» 

0 <E 
k 

O 

X 

•u cy 

10 Jt 

* s 

£ 5 

t c 

1 

R 

0 

K Cl 

/fair Ore 

thru 20 on 40 

0. 

14.10 

o. 

1.65 

L65 

4.37 

6.02 

27-4 

ffatv Ore 

thru do 

o. 

9.4o 

o. 

1.50 

l.5o 

4.18 

5.68 

26.5 

/J_ a 

thru 20 on 4o 

0.5 

5.28 

0.10 

3J8 

3.28 

2.81 

6.09 

S3. 9 

13- b 

thru 6o 

0.5 

2. IQ 

Loo 

4.10 

5.10 

1.06 

6./6 

82.8 

14- a 

thru 20 on 4 o 

t.O 

/ 45 

0.20 

4.63 

4.83 

0.78 

5.61 

86.1 

14- b 

thru 80 

1.0 

0.15 

0.10 

5.56 

5.6 6 

0.45 

6.11 

927 

' 15 - a 

thru 20 on 4o 

1.5 

o.ol 

0.10 

5.87 

5.97 

O.i 2 

6.09 

9Q.I 

15- b 

thru 8o 

1.5 

0.0 1 

O.IO 

5.41 

5.51 

0.5 Q 

6.09 

9o.a, 

16- a 

thru 2o on 4 0 

2.o 

o. 06 

0.12 

6.24 

6 36 

0. oz 

6.38 

99 a 

16 _ b 

thru 8o 

2.o 

o. os 

0.12 

5.36 

5.48 

0.61 

• 

6.09 

9 0. 1 

/7_ a 

thru 2o on 4o 

2.5 


0.1 X 

6.33 

6.45 

0.12 

6.57 

96.2 

1 7. b 

thru 8o 

2.5 


o to 

5.29 

5.39 

0.6 7 

6.06 

88.9 

/3- a 

thru 20 on 40 

3-0 

0.03 

o. tz 

6.36 

6. So 

0.13 

6.6 3 

98.2 

IQ _ b 

thru 80 

3.o 

0.0 6 

0.0 7 

5.22 

5.2$ 

0.75 

6.04 

87 7 


4 









































50 


ROBERT RHEA GOODRICH. 


Table- No. X — D 


H oas t/ny m the Gas 

Muffle Furnace at ten 

nper c 

rture 

a sc 

'•C., 

Sample. No. 

1 

N 

<5 

w 

tu 

*k 

3 

'O 

0 5 

N 

'v 

^ £ 

3 * 

0 i 

' o 

- *o 
s: S 

x> 

<3 C 
u 

c ^ 

<U \ 

,5 <s 

K S 

k. 

C 

U 

V 

V 
<u 

1 

'"o'"' 

0 

V 

X 

* 5 

5 

>sj **» 
1: 
C 

-k 

■5 S 

3*. 

O < 

s -S! 

>5 i 

O 

>. 

X 

$3 

3; <xj 

5 * 

N 

^ sc 

s^ 

s $ 

cu " 
<r 

-2 vs 

0 > 

k 

3 s 

VJ ^ 

V, 

0 

\ 

X 

3 w 

£ k 

SC ^ 

•u 

4! k 

-O C 

3 D 
k 0 

3 k 

k Hj 
V. £*. 

£ ' 
.0 3 
k k) 

’>■ 

0 

N 

X 

5 4! 

^ 2 
© ** 

* v 

3 0 
*»» 

s^* 

i * 

1 

3 

VJ 

3 c 

"Cl 4) 

3 5! 

3 qj 

* * 

X 

. q> 

3 k 
k) Q. 

5 

"> * 

0 k 
■k 0 

sc v 

'-k 

S 

:*s 

0 

k 

•> 

0 

k 

X 

t Q - 

« I 

«S 5 

3 s 

sc ... 

1 

C 

3 

j 5 

C 51 

Fare Ore 

thru 20 on 40 

o. 

14.10 

0. 

1.65 

1.65 

4.37 

6.02 


Ffdpy Ore 

thru do 

o. 

9.4 o 

0. 

1.50 

/. 5o 

4.16 

5.68 

26.5 

/9_ a 

thru 2o on 4o 

0.5 

2.64 

0.10 

3.94 

4-04 

/.SI 

5.55 

72. 9 

/9 _ b 

thru 8o 

0.5 

0.34 

0.07 

5.06 

5.15 

0.99 

6./4 

84.0 

2o. a 

thru 20 on 4-0 

to 

0.28 

0.0 7 

5.46 

5.53 

0.39 

5.92 

93.5 

>0 

o 

i 

o- 

thru 80 

/ .0 

0.03 

0.07 

4.96 

5.0 5 

0.95 

6.00 

84.3 

21.4 

thru 2o on 4o 

1.5 


O.07 

5.71 

5.78 

o.4l 

6.1 9 

93.4 

2/ _ b 

thru 80 

1.5 


0.05 

5.06 

5.13 

0.98 

6.11 

84.0 

2Z_ ^ 

thru 20 on 4o 

2.0 

0.02 

0.10 

6.16 

6.26 

0.49 

6.75 

92.6 

b 

thru 8o 

2.0 

0.03 

0.05 

4 94 

4.99 

1.07 

6.06 

82.3 

23-4 

thru 20 on 40 

2.5 


O.IO 

6.50 

6.60 

0.68 

7.28 

90. e 

2.3 _ b 

thru 80 

2.5 


0.05 

5.01 

5.06 

O.95 

6.0/ 

84.3 

2 4-. 4 

thru 20 on 40 

3.0 

0.0 A 

0.07 

6.10 

6.17 

0.62 

6.79 

9 0.9 

24-- b 

thru 80 

L 

3.0 

0.04 

0-C7 

4.87 

4.94 

1.01 

5.95 

83./ 















































TREATMENT OF COPPER ORES. 


51 


Table- No. Z^E 


< 

>5 

. x 

h 50 

X 

S X 

0 ^ 

* £ 

$ x 

* s. 

X <5 

V. 

50 

r * 

£ jc 

11' 

R ^ 

Perce n t 

Extra ct/on 

of 

Copper 

ffoastmg 
temperature 
500 °C. 

poasft ng 
temperature 
Goo 0 c . 

Nocist tng 
temperature 
725 0 C. 

fi ousting 
temperature 
850 0 C. 

«0 

0 ^ 

k. 

x X 

0 50 
x. 

0 ^ 

'X 

Xs 

50 

0 X 

Qo ^ 

^ * 

k 

-x 

50 

^ * 

X V 

0 K 
50 

0 <$ 
‘N ^ 

X 

C 

S 

50 

0 £ 

^ 50 

X- 
v ^ 

5 ^ 

c 

'K. 

~X 

50 

x r* 

Q> X 

<0 

0 T: 

cv X 

X 

k 

~X 

50 

n 

0 C 

X) <0 
-x 

? * 

x* 

'S 

<0 

$ 

0 

§ 5 

^ x 

to 

0 k 

w X 

X 

~x 

k 

50 

V 53 

0 x 

<0 <0 

X 

5 X 

tf. 

27-5 

26.5 

274 

26.5 

Z7.4 

26.5 

27.4 

26.5 

0.5 

<S9- 2 

93.3 

57.0 

74.4 

53.9 

82.6 

72.9 

84.0 

t.o 

SI. I 

Q7.8 

99.4 

92.2 

86. f 

92.7 

93 5 

84.3 

1.5 

93.3 

&&.4- 

98.4 

94.7 

981 

90.6 

93.4 

84.0 

2.0 

9o.S 

84.3 

98.3 

93.3 

99.8 

90. / 

9Z.8 

8Z.3 

2.5 

97-7 

92.4. 

/ 00.0 

93.3 

98.2 

68.9 

90.8 

84.3 

3.0 

95.8 

88.8 

97-5 

94.X 

98.X 

877 

9o.9 

83.1 


Note.: Data taken from Tables Nos. 2—A, 2-Bj 2—C y 
2— Dj and used tn the const ru ct ton of Cur/e Sheet No.2 

























52 


ROBERT RHEA GOODRICH. 


Table Mo. 3_A 


Ffoastinj 

m the Large 

6as Furnace 

at temperature 

6ot 

3 0 C. 

0 

<u 

5 

<5 

s 

1 

*>» 

V, 

qj * 

•v 

* 

$ ^ 

^ *> 

* 5 

V 

N 

V. 

C 

.* 

* * 

0 N 

^ V. 

' o 

- JO 

<V> ~ 

'O t 

*• 

-is. 

4) 

u 

V 

l 

S 

'"o'" 1 

O 

x 

* $ 

5 * 

|» 

v> 

4J 

4J £ 

•o ^ 

1J 

O Vi 

c: 

~ -v> 

E» i 

■> 

O 

X 

3 5 
1* 

M 

4 

s 

=3 ^ 

»o 

O 

■V. 

X 

n 

3 1 

■5 “V 
^ 0 

4! *. 

■0 s 

5 d 

VJ 

0 V. 

"3 t> 
<E 

5 J 

O 

■V. 

X 

^ •!> 

^ 5 

0 

^ V 

* 0 
^ "V" 

« 

Vo 

•s 3 

u 

^ ? 

0 V 

*v 

•> 

0 

•">% 

X 

^ $ 
55 

0 V 
•V <3 

■V’ ■<• 

1 V 

a c 
« 

55 

£ 

O 

X 

|s 

■v. a 

* 5 

* 5 

5 $ 

*3 

■V V 

1 

c 

0 

•* -v. 

vc 

if 

■*» qj 

■3' 

Ffatv Ore 

thru 

20 on 4-0 

o. 

u.to 

0. 

t.65 

1.65 

4.37 

6.02 

274 

B _ / 

•0 

O. 25 

13.10 

0.07 

t.67 

1.74 

4.32 

6.06 

28.7 

_ Z 

m 

0. 50 

H.20 

0.07 

2.34 

2.4 / 

4.04 

6.45 

37 4 

_ 3 

» 

»* 

0.74 

6.oo 

o.o 9 

2.46 

2.55 

4.29 

6.84 

37.3 

_ 4 

H 

/. 

S./6 

o.n 

3.60 

3-77 

2.97 

6.74 

55.9 

_ 5 

V 

/ .25 

3. IO 

0.26 

3.92 

4.16 

2-56 

6.74 

62.0 

_ 6 

n 

/ .50 

/. 70 

0.24 

5.04 

5.28 

1.60 

6.86 

76.8 

_ 7 

m 

■1.74 

1.50 

0.95 

4.75 

5.7o 

0.99 

6.69 

65.3 

_ 8 

V 

2. 

1.52 

1.38 

4.92 

6.30 

0.54 

6.84 

92.2 

- 9 

•* 

2.25 

/.Co 

/. 67 

4-61 

6.28 

O .60 

6.88 

91.4 

_ to 

m 

2.50 

1.72 

/• 72 

5.0 4 

6.76 

_ 

0.45 

7.21 

93.9 
































TREATMENT OF COFFER ORES. 


53 


Table No. 3-B 


Boasting m the Large Gas Furnace at temperature 600 “ C. 

O 

* 

V 

<5 

s 

1 

\ 

* 

C 

U 

1* 

<5 

*>> 

V ' 

0 $ 

<u 

N 

s 

'v 
s v. 

V. *> 

* ^ 

* < 

* * S 

$ {0 
s « 

>s Sj 

^ ts 

<u 

<3 c 
<0 

! * 

<*> \ 

I ^ 

Q 

<U 

O 

C 

<D 

1 

S 

'"'o'' 1 

O 

\ 

X 

1 « 

*1 
5 * 

v: s 

S! 

s s 

1 V 1 

^ 4“ 

v3 1 

r "o" 1 

0 

X 

$ 

^ V 

0 
$ <5 

N S 

^ c 

Q; S 

S 

^ * 

s ^ 

55 $ 

VO >3 

r 'o"' 1 

O 

■«» 

X 

^ 0? 
* g. 
s <5 

« * 

O. -K 

•0 ts 

3 <U 
\ ij 

0 V. 

H <U 
^ Q. 

53 1 

1 

.0 0 
K vj 

O 
— , 

X 

a « 

4! *. 
•5 5 

0 s 

H V 

5 0 

V 

$ * 

1 

0 

C> 

H) 

■< C 

<3 <0 

3 v! 

O 

X 

^ $ 
4 5 

0 V 

■u 

3$ 

C3 

■**. 

X 

>5 „ 

U -5! 

5 1 

'♦v ^ 

2 5 

3 v2 
* v 
AA 

5 

C; 

0 

^ -k 

s 5 
i 5 

V Bj 

Farr Ore 

Maw 40 cn 60 

0 . 

14.85 

0 . 

/.60 

1.60 

£.02 

6.62 

24.2 

B -// 

m 

0.25 

10.20 

0.19 

2.39 

2.56 

4.30 

6.88 

37.6 

_ /2 

m 

0.50 

6 .35 

0.41 

2.77 

3.18 

3.92 

7.10 

44.8 

_ 13 

* 

0.75 

4.06 

0.52 

4.08 

4.6 0 

2 . £8 

7 18 

64.1 

_ 14 

H 

/. 

1.84 

0.41 

£.63 

6. 04 

1.2 0 

7.24 

83.S 

_ IS 

N 

/ • 25 

1.04 

0.50 

6.31 

6.81 

0.37 

7/8 

94.9 

_ / 6 

m 

/ .50 

1. 15 

0.93 

5.96 

6.89 

0.23 

7/2 

96. 9 

- / 7 

M 

!.7S 

1.2 1 

1.00 

£.86 

6.86 

0.26 

7-12 

964 

_ /<? 

* 

2 . 

1.23 

I 17 

£.68 

6.85 

0.25 

7 . /o 

96 -6 

_ /5> 

• 

2.25 

1.26 

1.31 

£. 7/ 

7.02 

0. 

7 02 

too. 





i 

_1 

_ 





































54 


ROBERT RHEA GOODRICH 


Table A Jo. 3-C 


Ffoastmq in the. Larye 6as Furnace, at temper 

ature 

6oc 

> *C. 

o' 

* 

V 

$ 

S 

l 

<5 

C 

<D 

-k 

$ 

0 s 

D 

N 

'•s. 

\ 

si 

o ^ 

13 

^ ft 

<U „ 

^ £ 

4) 

£ ■* 

sb 0 

$ $ 

4 

C 

d) 

C 

<u 

i 

5 

0 

X 

V. 

»J y 

5 X 

* * 

5 ft 

■k 

$ c 

4) 

4) £ 

4) 

t *» 

'o V 1 
c; * 

-k 

<5 i 

■> 

>» 

X 

Vj 

? s’ 

5 ft 
v ' 

!: 

S>. 

$ $ 

4) * 
-< 4J 

* > 
0 > 

5 ^ 
Ei >5 

r ~o"'' 

0 

X 

5 4) 

* ^ 
'S 5 

s ft 
*4 

-S v 

J ° 

* 4. 
-0 0 

4. <J 

0 E 
ft 4) 
k Cc 

£ > 
£ vj 

r 'o'"> 

o 

4 . 

X 

a * 

4> X 

^ s 

o ^ 

>-> 4. 

c o 

>fc 

i * 

l 7^ J 

3 

E> 

•s^ 

X4 4) 

-5 K 

0 4) 

r 

■>» 

o 

k 

X 

ft **. 

-5 ^ 

+; -k 

vj « 

iS 

■> 

C> 

>» 

X 

* 4) 

v <: 

4» Q 

V» R 

-2 ^ 

5 c 

3 s 
^ 'o 

1 

cr 

ft 

^ ■+. 

5 S 
cs 

ffatv Ore 

thru So 

o. 

9.4c 

o. 

/. 5o 

1.50 

4. IQ 

5.66 

26.5 

B _ 2.0 


O. 25 

5.22 

0.55 

3.27 

3.82 

1 . 62 

5.64 

67-6 

- 2 / 

« 

0.50 

2.13 

0.96 

3.75 

4.7 / 

0.95 

5.66 

63.2 

_ 22 

- 

O.JS 

1.04- 

0.91 

4-28 

5.19 

0.47 

5.66 

91.6 

_ 23 

h 

/. 

1.20 

1.31 

3.87 

5.18 

0.46 

5.64 

920 

_ 24 

" 

/ .26 

/ .22 

1.53 

3.82 

5.35 

0.2 9 

5.64 

950 

_ 25 

•• 

1.50 

1.25 

1.48 

3.82 

5.3o 

0.32. 

5.6 Z 

94.4 

_ 26 

» 

/ .75 

J.32 

1.4 6 

3.62 

5.28 

0.34 

5.62 

940 

96.0 

- 27 

- 

Z. 

1.24 

1.41 

3.94 

5.35 

0.22 

5.57 














• 















































TREATMENT OF COPPER ORES. 


55 


Table No.3-P 


Ff oast 

inq m the Large Oas Furnace at temperature 6 oo°C. 

0 

<w 

5 

* 

1 

N 

<5 

\ 

C 

<u 

V 1/1 

0 * 

N 

S 

»o 

* 

V. 

V. 5J 

0 $ 

^ c 

•J 

^ <5 

50 r. 

^ C 

$ "< 

• > 

£ ^ 

s 

!U 

VJ 

C 

<0 

1 

O 

N 

X 

V. 

V) « 

1 ^ 

* 5 

' <S 

'V *-> 

^ V 

$ * 

<0 

<0 fcj 

1 S. 

<a 

* £ 
Vo * 

\ 

X 

$ 

3; <U 

5 «r 

N <0 

■U 

c: ^ 

5 g 

3l < 

X) 

O > 

5 "4 
VJ ^5 

■>> 

O 

X 

<u 

^ 1 
\ ^ 

* 0 

41 

-o c 

i 

^ LJ 

0 C 

V| nj 

-- Cl. 

5 ' 

0 

X 

3 * 

4 2- 
4 * 

^ O 

is 

cE 

1 

<o 

-X s 

"Q 4) 

<5 £ 

0 SO 

0 

-v 

X 

<D 

3 > - 
<o ^ 
5 

* * 

C ^ 

^ 0 

■*"3 ■< 

1 ^ 

V) ^ 

• s 

0 

K 

r ''o"' 1 

0 

X 

^ « 

V CE 

C q 
-** ^ 

V 

C 

1 

c 

0 

'S. V- 

^ c 
u m 

J t! 

(^) 

E ^ 

FfatY Ore 

Whole thru 2.0 

O. 

12.70 

o. 

/. 70 

/. 7o 

4.34 

6.04 

2 8.2 

-28 

* 

0.25 

10.24- 

0.12 

2.32 

2.44 

3. 70 

6.14 

39.3 

- 29 

H 

0.50 

7.56 

0.24 

2.6o 

2.8 4 

3.34 

6.18 

46-0 

_ 3o 

n 

O. 75 

5.16 

0.53 

3.3o 

3.83 

2-47 

6.3o 

60.8 

_ 31 

• 

• 

l . 

3.62 

0.5 7 

4.06 

4.63 

I .69 

6.32 

73.5 

_ 3Z 

it 

1 .25 

I. 4c 

0.5 / 

4.58 

5.09 

1.21 

6. 3o 

60.8 

- 33 

n 

/ .50 

o. 76 

0.41 

5.42 

5.83 

0.59 

6.4z 

9 0.8 

- 34 

m 

f. 75 

0.64 

0.48 

5.42 

5.9 0 

0.52 

6.42 

92.0 

_ 33 

n 

Z. 

0.84 

0.67 

5.35 

6.02 

0.26 

6.2 8 

96 0 

_ 36 

h 

2.25 

0.86 

0.62 

5.42 

6.04 

0.24 

6.28 

96.2 

-37 

to 

2.50 

0.72 

0.69 

5.35 

6.04 

0.24 

6.28 

96 2 

































56 


ROBERT RHEA GOODRICH. 


Table No.j-E 


fjoastmj m the Large Gas Furnace * 

it let 

nper* 

iture 

6 OC 

• C. 

Sample h/o. 

1 

* 

>■* 

V 

%> 

<5 

•! 

4) 

x 

<3 

< * 

8 5 

1 * 

5 * 
s $ 

N 

^ <5 

'Q C 

V 

C * 

!* 

V \> 

k v-> 

V 

C 

U 

c . 
<u 

1 

■> 

© 

•>>» 

X 

V 

-5! 4! 

* $ 

r* 

$ * 

' ? 
is. 

si 

s *» 
V3 k 

O 

>•* 

X 

V 

^ « 

5 <s 

J< 

Xi 

•fc. 

< (1 

-5 vs 

^ N 

V) >s 

O - 

X 

N £ 

S * 

V. "> 

^ s. 

S 0 

-5! V 

O © 

5 w 

^ O 

0 © 
s M 

•V. <k 
£ 1 
,0 S 
k ©> 

X 

-5 

0 * 
s V 

t 0 
\ 

c ■*»■ 

* * 

1 

-S c 

^ £ 

O <U 

5 ^ 
\ 

IP 

x. 

X 

5 -5? 

U 

■k 

0 V 
■*- 0 

■u 

1 V. 

^ c 

5$ 

;■* 

O 

X* 

X 

^ %) 

%j <: 

5 1 

5 s 

3 a 

Xi 

1 

0 

X ^ 

-K „ 

>5 ^ 

Ffa# Ore 

thru 20 on 40 

O 

14.10 

o. 

1.65 

1.65 

4.37 

6.02 

27.4 

A _ / 

m 

O. 2s 

IO ./8 

0.14 

2.27 

2.41 

3.59 

6.00 

40.2 

_ 5 

a 

/ .25 

2.30 

0.50 

4.11 

4.6/ 

1.6 7 

6.2 6 

73.5 

_ /o 

■ 

P..S 

0.8 9 

I.IO 

5. II 

6.21 

0.44 

6 - 65 

93.5 

F}aw Ore 

thru 4o on 80 

o. 

14.85 

O. 

1 60 

1.60 

5.02 

6.6 2 

24.2 

A _ // 

V 

O .25 

8.96 

0.19 

2.70 

2.89 

3.87 

6.76 

42 .a 

. IS 

4 

l .25 

0.94 

0.48 

62 / 

6.6 9 

0.31 

7-oc 

95.5 

- 19 

O 

2. 2 5 

0.69 

0.67 

6.19 

6.86 

0.09 

6.95 

968 

Flan Ore 

thru SO 

O. 

9.4o 

0. 

l.5o 

l.5o 

4.18 

5.68 

26.5 

A _ 20 


O. 25 

5.02 

o. 79 

3.49 

4.28 

1.38 

5*G>G> 

75.6 

_ 24 

* 

I .25 

0.71 

0.77 

456 

5.33 

0.33 

5.66 

94.3 

- 21 

m 

2. 

0.93 

0.96 

4.41 

5.37 

0.34 

5.71 

94.2 

Flan Ore 

whole thru 2o 

o. 

12.70 

O. 

1.70 

I.70 

4.34 

6.04 

28.2 

A _ 26 

• 

0.25 

3.76 

0.17 

2.27 

2.44 

3.81 

6.25 

39.1 

_ 32 

* 

1.25 

O. 93 

0. 55 

5.43 

5.98 

0.43 

6.41 

93.3 

_ 36 

• 

2.25 

o.87 

o. 72 

5.3o 

6.02 

0.22 

6.24 

96.5 


Note: This table, is the record of a check jer/es of roasts made, in 
the Large 6as Ffoast/ng Furnace __ results not plotted 






































TREATMENT OF COPPER ORES. 


57 


Table No. 3-F 


Comparison of total soluble Copper deter m / n e d by using dil. HC! 
with that determined by using dil. H X S 0* 

<u 

0 

'ts 

-k. 

>-> 

<5 

<a 

0 

4 

-9 

!• 

1 

■k 

<5 

° 5 

V 

N 

•k. 

<U 

s 

1) 

Q. 

1 

C) 

\ 

<5 

-k 

£ 

1.0 gram of sample 
bode d 20 mm. 
using ISO c.c. 
dil. HCi 

f 'OO c.c. cone, tfc/l 
/ diluted to lOOO C.C.J 

/. o gram of sample 
boiled 2o mm. 
using /So c.e. 
dil. (topercent) 
H x SO+ ' 

Total soluble 

Cu _percent 

i 

C 

5 - 

O V 

Q 0 

k Ik 

«M * 

Total soluble 
Cu — pe rcent 

1 

5 < 

* W 
s b 

*5 ^ 

A ~ to 

thru 2.0 on 4-0 

6.65 

6.21 ® 

93.5 

S .74 ® 

66.4 

- 19 

thru 4o on 60 

6.95 

6. & G> 

9 6- 8 

6.77 

97.4 

- 27 

thru do 

5.7 / 

5.37 

94.2 

5.45 

95.5 

_ 36 

whole thru 20 

624 

6.0-3. 

96.5 

5.9 2 

95.0 


* Work was repeated and remits checked 















Table No. 


58 


ROBERT RHEA GOODRICH 




* 

0 

s 

3? 

< 

Q 

V 

0 

V 
<D 

0 

V. 

c 

5 

h 

v 

\ 

v 

\ 

0 

c: 

N» 

0\ 

'C 

u 

<S 

N 

*?/ ~ P^ioaa4.X9 ny -7/ 
jad pauunsuo a (* r Of x H- tuza-tad 

6 / '£6 * ?usna&,99j l° IJ t 'A }° /'O 


O 

<X> 

On 

VO 

0 

K> 

vs 

vs 

00 

VS 

% 

<7/ ~pap30JtX3 

ny -qr Jad pa u/n suoo *0 S X H 

'S 

■'•I 

ON 

■^1 

<X) 

N 

<0 

00 

fO 

n 

h- 

Qo 

K> 

vS 

to 

• q/ ~ pacf d os/ 

3jo /0 U 04 . J3d p diM n r uoo *of x H 

vS 

? 

VS 

K 

vS 

0 

K 

n 

0 

VS 

*0 

tujojB ~ pa uf yoa/ aso 
/0 luoj 6 jad pau/nruoy * 0 S Z H 

CO 

O 

<0 

O 

0 . oee 

00 

VO 

’s) 

vo 

<0 

•>* 

0 

0 

KN 

cX 

0 

ISZO 

/uayjad 

~ b u 1 i/ a 0 a / t// pau/nsuoa *0 S z f-/ 

R 

O 

S 

O 

C) 

CO 

00 

o' 

»0 

0 

K) 

cx 

<o 

quayjad ~ BuiLfJoa/ 

ja+j-D uoij.n/or isr ('aajj.) * 0 S X H 

N 

ci 

oS 

N 

ON 

N 

On 

C< 

ON 

vo 

CO 

0 

K. 

KN 

K 

tuaojad ~ fuufooa/ 
ajoqaq u 0 / j.r >/0 r u / (aajj.J v o / r // 

o' 

>. 

* 

> 

•• 

■ 

* 

■ 

ay ~p 3 if o03/ qjo j.o u/ojB 
jad patn uoi^n/or fu/Lfypay 

O 

V. 


r 

* 

* 

* 

• 

/~_ a/duJO! ui nj im ~) quayjad 

/ °°'* pat aojt ra oj t*J - uo/q oojq *3 

(S 

*s 

V9 

ON 

vs 

ON 

N 

K 

cs 

<0 

Co 

* 

«3 

1X3 

aj 0 j.o u 01 a ad -<q/ 

~ paqyojqya nq) 

0 

* 

S3 

H 

c< 

On 

VO 

wi 

Os 

VS 

V9 

N 

o\ 

O 

0 

ON 

■v 

■*» 

VO 

!< 

*■« 

a/ du/p s j.o t** 1 q^oyjad ~nj 
1 ° ' ”3 !°* /°t°t V*t J a/qn/or /oqqi 

0 

CX 

vS 

<3} 

KN 

D 

4 

vS 

n 

Vo 

V0 

ON 

w> 

oo 

OD 

*0* 

f~_ a/ du/os qo qM "j /us?j 5 i/ 

1°°' * ny/ofOft^J ~ r >3 /o + oj 

<0 

\S 

VS 

« 

* 

* 

ft 

5 

c 

"O xaaj6ap ~ 6u/ qyoa / 

buunp p 3 a / 01 u/oiu 3Jntajadu/3j^ 

■>» 

fX 

* 

* 

> 

0 

0 

-V 

■ 

» 

fjnot/ 

~ 6 uil/?& 3 -j j.o uot/t?jnff 

v: 


vS 


*<3 

vs 

N 

if sauj 

~ / O / J 31 O U! J.O 3Z/f 

* 

$ 

0 

0 

$ 

$ 


* 

s 


* 

- 

‘ON z/uosr 



*1 

* 

*) 


N. 

ajo paqsooj to -o/d a/du/Pf 

O 

1 

4 


* 

• 

• 












































TREATMENT OF COPPER ORES 


59 


Ov 

nO 

to 

"N 

VS 

<3\ 

vS 

cy 

VO 

QO 


<3) 

vs 

<30 

wo 

<30 

<30 

VO 

o 

fs. 


1^ 

Ov 

to 

to 

<0 

y 

to 

"** 

o 

0\ 

ON 

ri 

»0 

N 

o 

o 

On 

0\ 

cn 


X) 


■'»« 

-*X 


N 


x5 

<k 





'r 

4 

cy 

ty 

s. 

■X. 

Ki 

•»)' 

y 

h* 

*> 

k> 


vS 

3 

0 

30 

<S 




vS 

VO 

CS 



cy 


0 

k) 

Vi 

00 


ts 


rx 

ri 

VS 

0\ 

Ov 


5 

to 

0 


On 

Ox 


<S 

0 

fs. 

o 




■*c 

cx 

to 

*5 



*< 

''i 

't 

't 

5 


" 


X) 

*0 

y 

vi 

vs 

v9 

vS 

o 

o 

y 

VS 

S? 

VS 

vS 

0 

0 

* 

vS 

vS 

vS 

VS 

o 

0 

xt 




**• 


vS 


Vo 

Un 

ifl 

VO 

K 

vs 


N 

ON 

On 

On 

K 

vs 



(k 

N. 

<N 

to 

y 

X3 


■*>* 



X) 


Xl 






T 

xi 

33 


CD 

CD 

Vo 

0 

N 

<30 

CO 

CO 

CO 


o 

N 

CO 

<0 

co 

<^) 


0 

ts 


o 

o 

S3 

CD 

»0 

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rs. 


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rs. 

<Xi 

NN 

»0 


Ov 

ON 

Ov 


to 



■>. 



"s 

CH 

CH 

<5 

o 

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o 


cs 

0< 


o 

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tx 

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vi 

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6 

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C) 

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to 

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ty 

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o 

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<y 

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d 

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o 


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(S 

ri 

CS 

k) 

■>. 

0 

►0 

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K 

rt 

a 


0 

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cy 

U) 

0 

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ON 

ON 

K 


cH 

rt 

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0 

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to 

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6o 


ROBERT RHEA GOODRICH. 


Table, Mo. 5 _ A 


£ lectrolytic. Copper F\efmmg P r a c t I ce 

f£o/. A to. 7 is based on Col No. 2 and Col No. l) 

In Ibis table are given Values of current density (Cd No. z) , and Voltage drop per 
cell (Col. hlo.3), record e d in respect iVe electrolytic copper refineries (Cot. No. t) 

The rest sta nee _ ohms _ (Cot. No. b) _ of an assumed miniature cell with a cathode 

Surface of 0 . 329 b sg.fl. (Col. No. A), ope r a t m y with Current value (Col.No.Sj, IS cal¬ 
culated so as to conform to the voltage drop (Cot No.3) When the miniature cell 

15 operated as a miniature copper refining cell with this r e s / s t a n ce _ ohm s —(Co/. 
No.bJ, and at current dens>ty (Co/. No 2) , it will be preduemy cathode copper (lb- 
per krv. hourj the same m amount as the respective re f i n e r les (Col No. / ) 

Seren American Electrolytic Coppe r Refineries 

• 

Actual 

values 

Pertaining to the 
miniature cell 

e 

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Column No. / 

Col. 

Ho. 2 

Col. 

No.3 

Col. 

No. 4 

Col. No 5 

Col. No. 6 

cor. 

N 0.7 

Fjariton Copper Co., Perth Amboy , N.J. 

ie 

0.30 

0.3 296 

5 .933 

0.0501 


American Smelting and Refining Co.,Mau rer. At. J. 

12.5 

0.30 

• 

4 .i2o 

0.0 7 27 


U. S. Metals ffefmmg Co., Chrome , N.J. 

2 0 

0.36 

• 

6.592 

O 0577 


Batbach Smelting and Refining Co., Newark, N.J. 

20 

0.30 

■■ 

6.592 

0.0 455 

0.062 

• 

Calumet and Heda Mining Co., Buffalo, N Y 

// 

0.22 

• 

3. 6,2.5 

0.06 07 


Boston and Montana Con. C. and S. M. Co.,G>reat Falls , Mont[ 

34 

0.60 

* 

II. 206 

0.0535 


Anaconda Copper Mining Co., An a c on d a , Mont 

IO 

0.30 

• 

3.2 96 

O.O^l O 



e Hof man _ A1 et a Hu r g y of Copper t p. 526 



























TREATMENT OF COPPER ORES 


6l 


Table- No. 5-0 


Electrolytic Copper Pe.fi/iiny Practice. 

In this 
surface 
Value ii 

commt 
kw. hou 
(Table 
refm m 

table the 
0.3296 sq 
this table 
’■cat cur 
r) repre 
S.A , Col. 
g ,/n Cur* 

(This table is based on Table No. S.A,Col No 7 ) 

resistance of the assumed miniature copper refining cell with cathode 
ft. IS taken as 0.062 ohms, arerage Value (Table S.A , Col A/ 0 . 7 ). The current 
is Varied so that current density names through the whole range of 
rent densities. The . corres pon d 1 nq cathode copper produced (lb. per 

sents the average practice m electrolytic Copper refineries enume rated 

No.tJ . This is plotted , “Copper deposited lb. per kw. hour e/ectroluhc 
'e Sheet No. S 

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Note : ‘ r Cu rre nt 


Hofmun _ Metallurgy of Copper 
d en i tty — a mpe res per sq.ft, 'always 


p. 4 98 , current eff/c iency 90 to 95 percent 
signifies per sg ft. of cathode surface 










































method of mtrodactna the 


ROBERT RHEA GOODRICH 


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TREATMENT OF COPPER ORES 


63 



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Table No. 5-C ( Continued ) 


64 


ROBERT RHEA GOODRICH 


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Table No. 


70 


ROBERT RHEA GOODRICH. 




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72 


ROBERT RHEA GOODRICH 


Table, No. 6_/4 


Preliminary tests to dete rrrnne the evaporation of the electrolyte 
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Mote : For elecfr tea t record of finishing Step t Tests No. 2 and A t See Curve Sheet 


TREATMENT OF COPPER ORES. 



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P. 

b 

-1 

r\ 

a 

rv p 
x x 

yi 

ib 

o 

£ 

k 

P 


(free). 

percent 

O 

p 

>4 

Vm 

JO 

£ 

vo 


Cu .percent 

351 


i 

Vm 

§ 


Hi S £>. (free). 

percent 

K> 


X 

KJ 

A/o. / Step and Mo. 2 Step 

Dura 

of / 

tton 
e st¬ 
ars 

a 

s 


Finishing Step 

ho 

> 

«b 

<0 

X 

<x 

fc 

x 

i 

Amperes 

1 

M 

J0 

1 

1 

Cathode surface per eeJt. 

sg. ft 

O 

X 

V-s 

X 

ft 

'b 

Kj 

0 

Kj 

0 

kj 

0 

Current density, 
amperes per sg. ft. 

X. 

Vm 

O 

X 

k 

O 

k 

o 

Average Vo/ts per cell 

k 

k 

k 

Number of celll m tenes 

VO 

VM 

8 

Vo 

& 

s'® 

B 

Theoretteat Cu deposited. 
g rams 

JO «0 

5s £ 

'o ^ 

*!| 

x<0 k 

£ 
"* 'S 

Actual Cu depos ited _ 
grams 

JO 

s 

P 

c 

JO 

fc 

Total M d. hours 


vo 

k 

K) 

Q 

>o 

.o 

§ 

Current efficiency.percent 

1 

X. 

S 

> 

Vm 

k 

> 

Vm 

KJ 

Cu dep.osi t eft . 

/b. per kw. hour 

-5- 

Vm 

-RT 

Sm 

2 

5 

NM 

M 

* 

Cathode surface per cell. 

sg. ft. 


VM 

vm 

vm 

Current density- 
amperes par sg. ft. 

0 

o 

XI 

O 

X 

Average vo/ts per cell 

' 

X 

X 

Number of celll m series 

0 

Kj 

V—S 

Kj 

Vm 

ib 

K> 

Vm 

k 

Kj 

V* 

k 

Theoretical Cu deposited. 

grams 

*n 

SC ^ 

o x 

*** 

_ p *> 
56 i 

« 

Actual Cu d epos / 1 ed 

grams 

ft 

'b 

JO 

i 

JO 

§ 

* 

» 

a 

s 

Total kw. hours 


>o 

y> 

QD 

x> 

Vm 

<m 

2 

P 

Current efficiency, percent 

N 

VM 

o 

X 

Ib 

Vm 

X 

k 

k 

Cu depos lied. 

/b. per km. hour 



Table No. 















































































74 


ROBERT RHEA GOODRICH 


Table No. 6 _ C 


, / Test No. a _ Lead Anodes (Anode No t) 

Finishing ep / Telt No , 4 . _ Cerlon Anodes dith sulphur dio 

L (Anode No.j) 

Mide. qat 


1 

»•* 

e 

4. 

VJ 

V 

"V 

Si 

V 

0 

rc 

? 3 
c 0 

r 

'O 

S 

O 

SI 

S 

£ 

if 

s* . 
- 

* 

Si 

* <V 
v 

* si 
s» v. 

V • 

V. X 

3 s 

^ 3 

*-s 

X 

0 

* * 
v . 

3 x 

4 * <1 

N 

v. . 
3 0 

£ 

1 

Si 

Vj 

3 

s* 

V. 

3 

« 

0 

% 

Vj 

vi 

V) 

1 

Si 

■4- 

■v. 

V» 

Si 

Si 

V 

0 

SJ 

£ 

3 

0 

3* 

4. 

c 

%) 

VJ 

V* 

Sj 

1 

«u 

■*- 

3> 

e 

■4- 

VI 

St 

*5 

s 

3 

VJ 

V* 

1 

■*4 

£ 
x 0 
-=? 2 

| 

1 

e 

4 * 

3* 

s 

•V. * 

V* c 

* 2 
c ^ 

3 

V) 

1 

1 

t 

c 

V 

3 

% 

-C 

Wn 

<r- ^ 

s $ 

"3 %N 

X 

*■ 

*% 

O' 

3. 

St. 

■3 

3 

V) 

? 

2 

— 

1 

s < 

•. 0 
•Xj VJ 

^ 1 

0 0 

•S ^ 

c 

~0 

V) 

V- 

1 

~S 

•i. 

0 

I" 

3 

Vj 

1 

X 

oo 

«2 

0 

•w 

Hk 

O 

4. 

5 

|V 

t • 

V -N 
-c. *0 

^ X 

»x ^ 

( > 

u 

ti 

a 

T 1 

* 

I 

^ s 

s < 

•< ^ 

Sw s 

Ji 

}3 

ri 

6 $ 
»: -s 

0 

0 

> 

X 

N -N 

^ 3 

|| 

1 

C 

Si 

0 4 . 

$ s 
- J 

•*> V. 

V, 

4- 

C 

Sj 

Vj 

V. 

V 

1 

S) 

>\ 

•0 

w 

VJ 

3 / 

3 

0 

"5 

■4~ 

0 

■4- 
'— 

0 

^r 

H 

1 ^ 

|3 

1 ^ 

v < 

A ^ 

0 < 
|-x 

3 Si 

3 ^ 
VJ ^ 


Col. 

Vo./ 

Col. 

Vo. a 

Col. 

No.J 

Col. 

No. 4 

Col. 

No. 5 

Col. 

No. 4 

Col. 

No . 7 

Col. 

No. 6 

Col. 

No. 9 

Col. 

NO. K 

Col. 

NO. II 

Col. 

Non. 

Col. 

NO. !3 

Col. 

No. !4 

Col. 
No. IS 

Col. 

No. 16 

—N 

V* 

4 

i 

s 

X 

H 

o 

3 

X 

% 

V 

1 

ri 

i 

%. 

H 

£ 

0 

*5 

1.6*6 

< 0)1 

5 no 

O .74 J 

39.6 0 







15.40 

1.60 

.1.55 

n 


- 



0X84 











3. 


• 



0X17 











4-5 


• 



0.J9I 











6. 




- 

0.594 











7.f 


• 



0.475 

24.70 

2.79o 

U.fo 

34.0 

2.925 


95 .2 




9- 


-■ 



0.4X5 











10.5 


-« 



0.399 











12. 


- 



o.joj 











1 35 


- 



0251 











IS. 


- 



0.206 

IO 60 

Ate ref* 

2780 

27-90 

72.0 

■ 


952 


1.71 

1.41 

1*5 


• 



0.141 









1.7* 


O 

•8. 


- 



O.IIJ 











<9 5 


• 



0.056 

Xox 

Jtenef 

2 593 

39.58 

92.2 

. 


887 


(2.13) 

I .09 

iff 


• 



0.049 

2.59 


34.09 

935 


1985 

( 795 ) 

15.80 

2.20 

0.94 

2o. 
















75- 

1 

! 

d 

1 

H 

O 

i 

V 

3 

C 

3 

S 

< 

0. 

2.5 

/X48 

0 6S91 

52 to 

0 . 74 J 

38 4o 







' 15.40 

0.75 


is 


- 



0.638 









• 


3 


■ 



0.591 









• 


4.5 


- 



0544 









• 


6. 


• 



O. 50 I 









• 


7-5 


• 


k 

0.453 









• 


9 


• 



0.401 









• 


to.jt 


- 



0.346 









• 


* 

** 

* 

0 

t 

n 


• 



O. 9 O 6 









• 


•*5 


• 



0.246 









■ 


X 

s 

0 

If. 


• 


d 

0.229 

11.90 

Atetoft 

2X7 

26jo 

* 9.1 

2.9X5 


914 


- 

3.18 


16.5 


* 



0.179 









- 


1 

■* 

18. 


• 



0.12 6 

6.55 

Atervye 

2.679 

32.05 

83-2 

- 


91.5 


• 

3.18 

I 

19-5 


• 



0.062 

X.27 

2.28 

3A.55 

88.9 

• 


78.0 


- 

2.72 

•K 

£ 

>9-85 


• 



0.073 

3.80 


34.80 

9o.x 


• 9 65 

(* 9 . 0 ) 

no 

• 

2.40 


20 . 

















o Copper dopeted firm to lh,s pent. Further electrolyse produces sponqu eeposit 

• Copper deposited firm end bright to this point. 

• On carrying the electrolysis to the So hour point .black copper rulph.de 
forms o* the, cathode. 








































































































TREATMENT OF COPPER ORES. 


75 


Example of Step Arrangement of Plant—Arrangement A. 


Copper Contents of the Electrolyte of the Steps held in 
Geometrical Ratio %. (except the Finishing Step). 



Feed- 

Copper 

Solution 

No. 1 
Step 

No. 2 
Step 

No. 3 
Step 

No. 4 
Step 

Finishing 

Step 

Relative copper contents 
of electrolyte. 

I 

X 

1 



T6 tO O 

Copper contents of elec¬ 
trolyte, percent. . 

6.0 

i -5 

0-375 



0.375 to 0 

H0SO4 contents of elec¬ 
trolyte, percent. 

0. 

6-93 

8.66 



8.22 tO 9.24 

Copper remaining in elec¬ 
trolyte, expressed as 
percent of that in feed. 


25 - 

6.25 



O. 

Copper deposited, ex¬ 
pressed as percent of 
that in feed. 


75 - 

18.75 



6.25 

Total copper deposited, 
expressed as percent of 
that in feed. 


75 - 

(Nos. 

1 and 2 
Steps) 

93-75 



(All Steps) 

IOO. 

Number of cells in series 
in each step. 


X 

X 

4 



X 

T2 

Relative electrode area 
per cell. 


1 

4 



8 

Relative C. D. in each 
step. 


1 

X 



X 


N. B.—When using carbon anodes with sulphur dioxide gas, the acid 
contents will be higher than given in these tables. 

These tables were worked up, assuming that there was no evaporation 
of the electrolyte during its passage through the plant, or what amounts to 
the same thing that water was added to the different steps equal in amount 
to the evaporation occurring in the different steps. 





















76 


ROBERT RHEA GOODRICH. 


Example of Step Arrangement of Plant—Arrangement B. 

Copper Contents of the Electrolyte of the Steps held in 
Geometrical Ratio Yz (except the Finishing Step) 



Feed- 

Copper 

Solution 

No. 1 
Step 

Relative copper contents 
of electrolyte. 

I 

l A 

Copper contents of elec¬ 
trolyte, percent. 

6.0 

2.0 

H 2 SO. t contents of elec¬ 
trolyte, percent. 


6.16 

Copper remaining in elec¬ 
trolyte, expressed as 
percent of that in feed. 


33-3 

Copper deposited, ex¬ 
pressed as percent of 
that in feed. 


66.7 

Total copper deposited, 
expressed as percent of 
that in feed. 


66.7 

Number of cells in series 
in each step. 


X 

Relative electrode area 
per cell. 


1 

Relative C. D. in each 
step. 


1 


No. 2 

No. 3 

No. 4 

Finishing 

Step 

Step 

Step 

Step 

1 

¥ 



I to 0 

O.667 



O.667 — O 

8.22 



8.22 tO 9.24 

II.I 



O. 

22.2 



1 1.1 

(Nos. 

1 and 3 



• 

Steps) 



(All Steps) 

88.9 



IOO. 

X 



X 

3 • 



¥ 

3 



6 

H 



Ye 



























TREATMENT OF COPPER ORES. 


77 


t 


Example of Step Arrangement of Plant—Arrangement C. 

Copper Contents of the Electrolyte of the Steps held in 
Geometrical Ratio (except the Finishing Step). 



Feed- 

Copper 

Solution 

No. 1 
Step 

No. 2 
Step 

No. 3 
Step 

No. 4 
Step 

Finishing 

Step 

Relative copper contents 
of electrolyte. 

I 

X 

X 

X 

l 

TS 

tV to 0 

Copper contents of elec¬ 
trolyte, percent. 

6.0 

3-0 

i-5 

0.75 

0-375 

0.375 to 0 

H2SO4 contents of elec¬ 
trolyte, percent. 


4.62 

6-93 

8.08 

8.66 

8.66 to 9.24 

Copper remaining in elec¬ 
trolyte, expressed as 
percent of that in feed. 


50 

25 

12.5 

6.25 

0 

Copper deposited, ex¬ 

pressed as percent of 
that in feed. 


50 

25 

12.5 

6.25 

6.25 

Total copper deposited, 
expressed as percent of 
that in feed. 


50 

(Nos. 

1 and 2 
Steps) 

75 

(Nos. i, 

2 and 3 
Steps) 

87-5 

(Nos. 1, 2, 
3 and 4 
Steps) 

93-75 

(All Steps) 

IOO. 

Number of cells in series 
in each step. 


X 

X 

2 

X 

4 

X 

8 

X 

8 

Relative electrode area 
per cell. 


I 

2 

4 

8 

l6 

Relative C. D. in each 
step. 


I 

X 

X 

X 

1 

T 6 























78 


ROBERT RHEA GOODRICH. 


Example of Step Arrangement of Plant—Arrangement D. 


Copper Contents of the Electrolyte of the Steps held in Arithmetical 
Difference, 1.5 percent (except the Finishing Step). 



Peed- 

Copper 

Solution 

No. 1 
Step 

No. 2 
Step 

No. 3 
Step 

No. 4 
Step 

Finishing 

Step 

Relative copper contents 
of electrolyte. 







Copper contents of elec¬ 
trolyte, percent. 

6.0 

4-5 

3-0 

1-5 


1.5 to 0 

H2SO4 contents of elec¬ 
trolyte. percent. 


2.31 

4.62 

6-93 


6.93 to 9.24 

Copper remaining in elec¬ 
trolyte. expressed as 
percent of that in feed. 


75 

50 

25 


O 

Copper deposited, ex¬ 
pressed as percent of 
that in feed. 


25 

25 

25 


25 

Total copper deposited, 
expressed as percent of 
that in feed. 


25 

(Nos. 

1 and 2 
Steps) 

50 

(Nos. 1, 

2 and 3 
Steps) 

75 


(All Steps) 

IOO. 

Number of cells in series 
in each step. 


X 

X 

X 


X 

Relative electrode area 
per cell. 


1 

i -5 

3 


6 

Relative C. D. in each 
step. 


1 

% 

A 


>4 































Curve Sheet No. 


TREATMENT OF COPPER ORES. 


79 


Cumutert / f'e percent 


>o 

o 


o 


0\ 

0 


0 


H 

^ o 


s 

fc. 

0 


0 


Xi 

o 


o 




o 


o 


Nl 

<6 


% 

o 


Vo 

o 


\ 

<4 







TA 

r U 2 

o or, 

4a 

—/Dt. 

sh 

por 

f/Ot 

7 








4S.. 

4 fit 

'rce 

n t 

of 

who 

te 








O 

v 
















■v ; 

L 5%. . 

r hru 

4o 

on 

80- 

mes 

b pt 

irttc 

' n 







\va 

\*v 

2. 

2.9 / 

ere 

ent 

of 

wh < 

>/e 








\"3 
















,C4 

A* 3 

- 














• 















\o» 

\f\ 
\-v 















\ ^ 

1* 

\\ 

> 















r\ 

> 

f6 















\v* 









cv 

>, 





\ Tt 

'iru 

8o - 

mes 

b pi 

r\ 

in 



>D 

-Ai 

r»j 





| 

M.7 / 

>erct 

3 .nt < 

of kV 

ho/t 

> 




* 















K- 

* 














bt> 

< 













L R- 

u 











































0 


s 

06 

O 


h 

0 

o 






















































8o 


ROBERT RHEA GOODRICH. 



Curve Sheet No. 2. 













































Percent Coppe.r 


TREATMENT OF COPPER ORES. 


8 l 



6 


Percent S u / p 
























































Percent Copper 


82 


ROBERT RHEA GOODRICH. 



J u / p h 































































Percent Copper 


TREATMENT OF COPPER ORES. 


83 












































































8 4 


ROBERT RHEA GOODRICH. 



Curve Sheet No. 3-D. 


Percent Sul 




















































Pounds Oil of // trio/ (€€* BuumeJ consumed per pound Copper ex fracfed 


TREATMENT OF COPPER ORES. 


85 



3 


Curve Sheet No. 4. 




















































































86 


ROBERT RHEA GOODRICH. 
























































Curve Sheet No. 6. 


TREATMENT OF COPPER ORES. 


8 - 


Percent _ Copper- a/epos/ted _ Current efficiency 


















































































88 


ROBERT RHEA GOODRICH. 


Os 

c; 



















































TREATMENT of copper ores. 


89 



fo/utton 1 to /each ore 













































90 


ROBERT RHEA GOODRICH, 






















































TREATMENT OF COPPER ORES 


91 



jo/utton to /each ere 





























































92 


ROBERT RHEA GOODRICH. 


VITA. 

Robert Rhea Goodrich. 

Apr., 1864—Born, Hartford, Conn. 

June, 1885—Graduated B. S. in Mining Engineering, Mass. Inst. Tech. 

1885-1890—Mining engineer for principal collieries in the New River Coal 
Fields on the Chesapeake and Ohio Railway; (1886) member 
American Institute of Mining Engineers. . 

1890-1893—Mining engineer for most of the new collieries on Elkhorn 
Creek, in the Pocahontas Coal Field on Norfolk & Western 
Railway; also (1893) manager of Elk Ridge Colliery. 

1893-1895—Surveyed and put in irrigation plants for improvement of 
property, El Paso. 

1895- 1896—Assayer and chemist for Chihuahua Mining Company, Chi¬ 

huahua, Mexico; traveled through Mexico and Northwest, vis¬ 
iting mines and smelters. 

1896- 1898—Mining engineer and assistant superintendent Helena Mining 

Company, mines of Chihuahua Mining Company; (1898) 
superintendent for the company. 

1899-1902—Student “M. I. T.”: 1901, graduated B. S. in Mechanical En¬ 
gineering; 1902, degree M. S., majoring in Electrical Engi¬ 
neering. 

1902-1904—“Engineering Apprentice,” Westinghouse Electric Manufactur¬ 
ing Company, East Pittsburgh, Pa.; (1903) Member American 
Society of Mechanical Engineers. 

1905-1906—Boston and Montana Consolidated Copper and Silver Mining 
Company, Great Falls, Montana: smelterman; draughtsman; 
studied concentration. 

1907- —Visited and studied important mining sections, principally 

Coeur d’Alene, Idaho, and Butte, Montana, districts. 
Superintendent Columbus Borax Company mine, Lebec, Kern 
County, California; installed roaster. 

1907-1914—Professor of Metallurgy, charge of Bureau of Mines and As¬ 
saying, University of Arizona; studied Arizona copper fields; 
limited research work for Arizona mining companies; visited 
important mining sections of United States. 

1913-1914—Sabbatical year at Columbia University; (1914) member 
American Electrochemical Society. 











































































































































































. 

















. 























. 


Ill 


























































































